•••^•M 


COPPER  REFINING 


°MG  Qraw-MlBook  &  1m 

PUBLISHERS     OF     BOOKS      F  O  R^ 

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Chemical  3   Metallurgical  Engineering 

Electrical  Merchandising 


COPPER  REFINING 


BY 

LAWRENCE  ADDICKS 

CONSULTING    ENGINEER,  *NEW    YORK  CITY 


FIRST  EDITION 


McGRAW-HILL  BOOK  COMPANY,  INC. 

NEW  YORK:    370  SEVENTH  AVENUE 

LONDON:    6  <fe  8  BOUVERIE  ST.,  E.  C.  4 

1921 


COPYRIGHT,  1921,  BY  THE 
McGRAW-HiLL  BOOK  COMPANY,  INC. 


THE     MAPJ.E     I'KI.SS     YOHK     PA 


To 


435338 


PREFACE 

Electrolytic  copper  refining  was  for  so  many  years 
conducted  under  conditions  of  strict  commercial  secrecy 
that  but  little  has  been  published  regarding  the  principles 
of  operation,  as  distinct  from  descriptions  of  individual 
plants. 

This  little  book  comprises  a  series  of  articles,  each  dealing 
with  one  of  the  problems  of  refining,  which  originally 
appeared  in  Chemical  and  Metallurgical  Engineering. 

Whatever  of  value  it  may  contain  is  due  to  the  fact 
that  it  is  almost  entirely  a  record  of  personal  experience  in 
a  field  where  but  little  literature  is  available. 

LAWRENCE  ADDICKS. 

NEW  YORK, 
December,  1920. 


vit 


CONTENTS 

PAGE 

PREFACE vii 

CHAPTER 

I.  METAL  LOSSES 1 

II.  METALS  IN  PROCESS 18 

III.  TANK  RESISTANCE 38 

IV.  CURRENT  DENSITY 54 

V.  CURRENT  EFFICIENCY 71 

VI.  IMPURITIES 80 

VII.  BY-PRODUCTS 100 

VIII.  FURNACE  REFINING 124 

IX.  THE  REQUIREMENTS  OF  REFINED  COPPER 143 

X.  COPPER  FROM  SECONDARY  MATERIAL 158 

XI.  THE  POWER  PROBLEM 171 

XII.  ELEMENTS  OF  DESIGN 183 

XIII.  APPLICATION  TO  OTHER  FIELDS 194 

INDEX.                                                                                                .  203 


COPPER  REFINING 


CHAPTER  I 
METAL  LOSSES 

It  is  only  recently  that  the  general  question  of  metal  losses 
in  metallurgical  practice  has  begun  to  receive  anything  like 
the  attention  it  deserves.  In  any  scheme  of  metallurgical 
treatment  it  is  always  easy  to  recover  say  50  per  cent  of 


Pis 


Scales 

Pig 

riuxes 


Market  Shapes       By-Prcdncts 
Copper 


Slag 


Silver 


Gold 


FIG.  2. — Diagram  of  process. 


the  value  in  an  ore  by  some  simple  direct  process.  Then 
improvements  are  made  which  will  bring  the  total  recovery 
up  to  perhaps  65  per  cent,  but  this  additional  15  per  cent 
has  cost  more  per  unit  than  the  first  50  units.  The  proc- 
ess gradually  develops  in  this  way  until  the  additional 
complexity  brings  the  operating  cost  up  to  a  point  where 


3\ :  i* '  ' '  COPPER  REFINING 

further  recovery  would  cost  more  than  the  value  of  the 
additional  metal  recovered.  It  is  also  soon  found  that  a 
high  recovery  process  can  profitably  treat  a  leaner  ore,  and 
this  adds  another  variable  to  the  most  economical  effi- 
ciency equation. 

Nor  is  the  study  of  the  various  sources  of  metal  loss  a 
simple  analysis  of  slag  and  stack  losses,  although  these 
are  among  the  largest  and  most  obvious  sources.  While 
we  are  now  considering  copper  refining,  or  the  separation 
of  blister  copper  into  fine  copper,  gold  and  silver,  the 
method  of  examination  will  apply  equally  well  to  any  other 
metallurgical  process. 

Figure  1  gives  a  flow  sheet  of  the  process  under  con- 
sideration and  in  Fig.  2  this  flow  sheet  is  reduced  to  its 
simplest  terms.  The  sources  of  loss  to  be  considered  may 
be  classified  as  follows: 

A.  Weighing: 

a.  Incoming  blister  copper. 

b.  Outgoing    copper. 

c.  Outgoing  silver  and  gold. 

B.  Sampling: 

a.  Moisture. 

b.  Errors  in  method. 

c.  Salting. 

C.  Assaying: 

a.  Assay  methods. 

b.  Splitting  limits. 

c.  Assay  errors. 

D.  Slags: 

Cupola  slag. 

E.  Stack  Losses: 

a.  Anode  furnace  stacks. 

b.  Refining  furnace  stacks. 

c.  Silver  refinery  stacks. 

d.  Cupola  stack. 

F.  Process  Lossess 

a.  Silver  and  gold  in  outgoing  copper. 

b.  Silver  in  outgoing  gold. 

c.  Gold  in  outgoing  silver. 

d.  Minor  losses. 

e.  Process  margins. 


METAL  LOSSES 


G.   Handling  Losses: 

a.  Wind  losses. 

b.  Theft. 

c.  Solution  losses. 

d.  Soil  losses. 

e.  Slimes  losses. 


A.  WEIGHING 


Scales  are  about  the  most  precise  pieces  of  physical 
apparatus  we  have.  In  the  laboratory  when  we  get  a 
sensitiveness  of  0.01  mg.  with  a  capacity  of  1  gram,  we 
are  working  to  0.001  per  cent  and  even  lesser  ratios  are 
obtainable.  In  large-capacity  scales  the  best  precision 
obtainable  is  about  0.01  per  cent;  that  is,  if  we  want  to 
be  sure  of  the  nearest  pound  we  must  not  use  drafts 
exceeding  10,000  Ib.  Using  a  200,000-lb.  railroad  scale 
which  to-day  can  be  made  distinctly  responsive  to  a  20- 
Ib.  change  in  load  gives  the  same  precision,  but  does  not 
equally  satisfy  a  customer  who  places  great  stress  on  in- 
dividual pounds. 

It  is,  therefore,  customary  to  weigh  outgoing  copper  in 
5-ton  drafts  under  such  conditions  that  the  refineries 
have  been  able  to  stand  upon  their  weights  as  final,  pro- 
vided the  consignee  agrees  with  the  count  of  pieces.  In 
order  to  accomplish  this,  two  platform  scales  are  placed  in 
tandem,  separate  weighers  taking  readings  and  comparing 
figures  before  passing  a  lot.  These  scales  are  either  of  the 
overhung  type  where  the  knife  edges  are  overhead  and 
always  in  sight,  or  baby  railroad  scales.  The  latter  are  to 
be  preferred  as  they  are  free  from  the  obstructions  above 
the  floor  line  and  it  is  possible  to  arrange  the  live  platform 
so  that  an  oncoming  car  divides  the  shock  between  two 
pairs  of  knife  edges,  greatly  saving  wear. 

The  standard  of  weight  is  furnished  by  the  certification 
of  the  U.  S.  Bureau  of  Standards  of  a  50-lb.  brass  weight  as 
a  "  Class  B  or  Working  Standard."  Against  this  are 
checked  a  sufficient  number  of  50-lb.  cast  iron  weights, 
adjustable  by  lead  shot,  to  make  up  a  full-scale  load. 


4  COPPER  REFINING 

These  weights  are  not  suitable  for  every-day  use  as  they 
afford  a  great  surface  for  the  condensation  of  moisture, 
collection  of  dirt,  etc.,  so  a  test  car  loaded  with  very  heavy 
cast  iron  weights  is  checked  against  them  every  two  or 
three  weeks  and  this  test  car  is  used  for  daily  checks  on 
the  merchant  scales.  In  this  way  it  is  quite  practicable 
to  maintain  the  desired  precision  of  1  Ib.  in  10,000.  As 
identical  methods  are  used  for  weighing  incoming  blister 
and  outgoing  wirebars  there  is  but  small  chance  of  tracing 
any  copper  losses  to  scales  in  a  modern  plant. 

In  weighing  the  silver  and  gold  the  situation  is  somewhat 
different.  The  incoming  silver  and  gold  are  weighed  as 
blister  copper  as  outlined  above,  while  the  outgoing  ship- 
ments are  weighed  as  bullion.  As  the  gold  is  always 
shipped  to  a  IT.  S.  Assay  Office  whose  weights  are  final,  we 
again  come  to  government  standards  and  have  automati- 
cally a  weekly  check  against  the  refinery  bullion  balance. 
In  one  of  the  large  refineries  a  year's  returns  showed  a 
difference  of  but  0.0015  per  cent.  As  a  bullion  balance 
is  furnished  with  a  set  of  weights  instead  of  a  sliding  poise, 
it  is  comparatively  easy  to  make  mistakes  in  reading.  To 
lessen  this  danger  it  is  well  to  provide  a  direct-reading 
spring  balance  upon  which  a  check  to  the  nearest  ounce 
can  rapidly  be  made. 

One  of  the  most  fruitful  sources  of  error  in  weighing  lies 
in  the  inaccurate  taring  of  cars.  Slight  differences  are 
bound  to  occur  daily  as  the  oil  works  out  of  the  bearings 
but  this  is  not  serious.  Wear  and  tear  gradually  lowers 
the  weight  of  a  car  and  it  should  be  tared  at  least  every  ten 
days,  the  weight  and  date  being  painted  on  each  side  of  the 
car  and  the  rule  adopted  not  to  use  any  out-of-date  cars. 
Similarly  a  car  sent  to  the  shop  for  repairs  should  have  the 
weight  painted  out. 

B.  SAMPLING 

a.  Moisture. — It  is  generally  assumed  that  copper  carries 
no  moisture.  This  question  was  first  raised  a  good  many 
years  ago  when  some  blister  copper  from  Australia  which 


METAL  LOSSES  5 

had  been  more  or  less  immersed  in  bilge  water  in  transit  was 
found  to  carry  about  0.5  per  cent  of  moisture.  Pig  copper 
is  quite  porous  and  a  quart  of  water  poured  on  the  face  of  a 
pig  will  vanish  as  if  absorbed  by  blotting  paper.  The 
same  question  arises  when  smelters  quench  converter  bars 
in  boshes.  About  all  that  can  be  done  in  such  cases  is  to 
make  careful  tests  in  a  drying  oven  on  certain  pigs  and 
apply  the  correction  found  to  the  lots  represented.  It  is 
evident  that  a  very  serious  source  of  loss  may  pass  unsus- 
pected if  this  is  not  watched,  as  the  refinery  will  be  account- 
ing for  water  as  copper,  silver  and  gold. 

Then  we  have  the  question  of  weighing  material  that 
has  been  exposed  to  inclement  weather.  This  should 
never  be  done,  but  sometimes  it  has  to  be  done  in  order  to 
save  large  delays,  and  it  is  then  necessary  to  impose  an 
arbitrary  allowance  by  agreement.  The  weighing  of  hot 
wirebars  does  not  seem  to  cause  any  appreciable  error,  the 
rising  air  currents  being  negligible.  Weighing  in  the 
wind  is  more  serious  and  scales  should  always  be  well 
protected. 

b.  Errors  in  Sampling  Method. — The  silver  and  gold 
contents  in  a  ton  of  blister  copper  may  be  worth  anywhere 
from  $10  to  $500,  although,  fortunately  for  the  refiner's 
peace  of  mind,  there  are  very  few  very  rich  bullions.  One 
hundred  dollars  per  ton  would  be  an  ordinary  figure, 
however.  It  can  readily  be  seen  that  a  small  error  in 
determining  the  true  metal  contents  of  blister  copper 
would  have  an  alarming  effect  upon  the  refiner's  profits. 
A  great  deal  of  experimental  work  has  therefore  been  done 
in  developing  standard  sampling  methods. 

Blister  copper  usually  runs  98  per  cent  to  99  per  cent 
copper,  the  bessemerizing  of  matte  having  reduced  the 
impurities  to  small  values.  The  molten  blister  is  poured 
into  open  molds  and  whether  or  not  the  copper  is  after- 
wards quenched  in  water  it  is  always  allowed  to  set  before 
being  removed  from  the  mold.  The  pig  therefore  cools 
most  rapidly  on  the  sides  and  bottom,  the  heat  being  ab- 
sorbed by  the  metal  mold,  and  the  last  spot  to  cool  is  the 


6  COPPER  REFINING 

upper  central  portion.  The  silver  and  gold  and  other 
impurities  form  a  complicated  series  of  solid  solutions  and 
eutectics  with  the  copper,  the  richest  portion  in  precious- 
metal  values  being  the  last  to  solidify. 

It  is  evident  that  any  system  of  sampling  must  bear 
this  in  mind.  Theoretically  this  would  require  that  when 
a  sampling  templet  is  laid  out  with  a  series  of  holes  for 
drilling,  these  should  be  arranged  to  straddle  and  not  pierce 
the  center  line  of  the  pig,  as  no  compensating  holes  can  be 
drilled  in  the  absolute  edge  of  the  piece.  Practically  this 
does  not  seem  of  great  importance  as  experiments  have 
shown  that  the  richest  zone  often  lies  in  a  ring  around  the 
center  rather  than  in  the  actual  center,  but  it  is  wise  to 
follow  the  theoretically  proper  practice  as  different  com- 
binations of  impurities  produce  quite  different  results. 

The  old-style  pig,  shaped  more  or  less  like  a  loaf  of  bread, 
is  now  no  longer  in  use  except  in  South  America  where 
purchases  of  bullion  are  generally  made  by  lot  at  auction 
rather  than  by  contract.  These  old  pigs  were  6  in.  thick 
and  the  segregation  defied  accurate  sampling  at  a  reasonable 
cost.  The  refineries  have  succeeded  in  introducing  a  bar 
about  18  by  27  in.,  2>^  or  3  in.  thick,  weighing  about  300 
lb.,  with  just  sufficient  draft  to  facilitate  removal  from  the 
mold  and  preferably  unquenched.  This  readily  yields  an 
accurate  sample,  stacks  well  and  is  heavy  enough  to  be  an 
economical  unit  in  handling. 

The  Japanese  still  make  very  small  pigs  weighing 
some  50  lb.  each.  These  cause  heavy  sampling  and  hand- 
ling costs  and  are  therefore  undesirable.  Once  in  a  long 
while  copper  comes  from  some  out-of-the-way  corner  of 
the  world  which  has  been  cast  in  pots  and  these  cases  call 
for  considerable  knowledge  and  ingenuity  on  the  part  of 
those  in  charge  of  the  sampling. 

A  number  of  other  precautions  must  be  taken  to  secure 
an  accurate  sample  of  pig  copper.  There  must  be  a 
sufficient  number  of  holes  in  the  templet — that  is,  one  hole 
must  not  represent  too  large  an  area,  otherwise  experience 
has  shown  that  the  sample  will  be  too  high.  In  fact,  care- 


METAL  LOSSES  7 

less  sampling  will  generally  give  results  on  the  high  side. 
On  the  standard  pig  a  quarter  templet  with  forty-eight 
>^-in.  holes  is  generally  used;  this  gives  an  area  of  about  2>^ 
sq.  in.  per  hole.  Each  hole  is  in  the  center  of  a  rectangle, 
arranged  so  that  forty-eight  such  rectangles  would  just 
cover  one-quarter  of  the  area  of  a  pig.  The  hole  is  drilled 
completely  through  the  pig,  each  successive  pig  sample 
being  marked  at  the  next  hole  in  rotation  on  the  templet. 
In  these  days  of  large  converter  pours  it  is  seldom  necessary 
to  drill  every  pig,  one  in  three,  four  or  even  five  being 
sufficient. 

It  must  be  realized  that  there  is  always  more  or  less  dirt 
on  the  surface  of  a  pig  and  also  that  the  set  side  has  an 
oxide  film  which  may  run  quite  different  in  values  from  the 
body  of  the  metal.  It  has  been  proved  that  decidedly 
different  samples  are  obtained  from  a  lot  depending  upon 
whether  the  pigs  were  drilled  face  up  or  face  down,  and 
further  that  while  for  a  given  brand  of  pig  these  differences 
are  generally  consistently  in  one  direction,  some  brands 
would  be  the  reverse  of  others.  This  dilemma  has  been 
met  by  using  the  so-called  top  and  bottom  method.  Origi- 
nally every  other  pig  was  turned  bottom  up,  but  it  was  found 
equally  satisfactory  to  drill  every  other  lot  bottom  up. 
No  watei  or  lubricant  should  be  used  on  the  drill  as  this  will 
tend  to  oxidize  the  hot  drillings. 

A  splash  sample  taken  at  the  smelter  by  collecting  a 
fraction  of  the  converter  pour  as  shot  in  water  will  always 
show  high  metal  values  as  against  the  refinery  as  such  a 
sample  is  automatically  freed  of  dirt  and  slag.  For  a  long 
time  the  smelters  felt  that  their  splash  samples  were  correct 
and  that  the  refineries'  samples  low,  but  it  has  been  shown 
that  a  properly  taken  sample  at  the  converter,  using  no 
water,  can  be  made  to  check  drill  samples  on  the  pig. 

Enough  has  been  said  to  show  that  sampling  methods 
are  a  matter  of  the  highest  importance  in  any  study  of 
the  sources  of  metal  loss,  as  the  refiner  has  to  pay  on  the 
values  shown  by  the  sample  whether  or  not  they  exist  in 
the  pig. 


8  COPPER  REFINING 

c.  Salting. — The  contamination  of  a  sample  during 
preparation  may  cause  serious  errors  and  consequent  ap- 
parent metal  losses.  Nowadays  this  is  seldom  done  de- 
liberately, although  the  large  laboratories  generally  have 
one  or  two  interesting  stock  tales  of  this  character.  It 
may  easily  be  done  unconsciously,  however,  if  silver- 
refinery  slags,  dore  samples,  etc.,  are  handled  in  the  same 
department  as  incoming  pig  copper.  The  only  sure  way 
is  absolutely  to  separate  such  different  classes  of  work. 

C.  ASSAYING 

a.  Assay  Methods. — In  the  early  days  an  assay  method 
was  supposed  to  represent  the  process  to  be  used  and  the 
assay  value  was  admittedly  lower  than  the  true  contents 
by  an  amount  approximately  equal  to  the  expected  loss  in 
treatment.  In  other  words,  the  assay  value  was  the 
recoverable  value.  To-day  the  tendency  is  to  adopt 
methods  which  will  give  the  highest  result  compatible 
with  reliability  but  which  still  do  not  in  all  cases  give  full 
contents.  The  electrolytic  assay  for  copper  does  give 
full  value;  in  fact  the  chief  danger  in  the  handling  of  this 
method  is  that  some  impurities  may  be  deposited  on  the 
cathode  and  weighed  up  as  copper,  to  the  disadvantage 
of  the  refiner. 

The  so-called  uncorrected  combination  assay  for  silver 
in  which  the  bullion  is  dissolved  in  nitric  acid,  the  silver 
precipitated  as  chloride  which  in  turn  is  scorified  to  gather 
the  silver  (and  gold)  into  a  lead  button,  the  button  being 
cupeled  and  the  resultant  bead  weighed  as  silver  plus  gold, 
does  not  give  full  value  of  silver  by  from  2  to  3  per  cent. 
The  loss  is  due  to  values  in  scorifier  slags,  to  cupel  absorp- 
tion and  to  volatilization.  The  main  argument  for  this 
method  is  that  when  conscientiously  run  it  will  duplicate 
results  very  closely  while  there  is  at  present  no  really 
satisfactory  method  of  determining  the  absolute  silver 
contents  of  foul  bullion. 

Skill  is  required  in  maintaining  the  proper  temperature 
when  cupeling;  it  must  be  hot  enough  to  make  the  button 


METAL  LOSSES  9 

"blick"  at  the  finish,  or  it  is  not  pure  and  other  elements 
are  weighed  as  silver;  on  the  other  hand  it  must  be  cool 
enough  to  show  "feather"  litharge  around  the  edge  of  the 
cupel  when  finished,  or  the  volatilization  loss  will  be 
excessive. 

Mismanagement  in  the  laboratory  may  easily  aifect 
the  apparent  or  financial  metal  losses.  In  some  cases 
the  " corrected"  assay  is  used,  where  the  losses  in  scorifier 
slags  and  cupels  are  determined  and  added,  bringing  the 
recovery  up  to  about  99  per  cent.  Nothing  is  gained  in 
accuracy  and  when  once  the  particular  bullion  is  known,  the 
amounts  of  these  losses  are  very  constant  and  can  just  as 
well  be  allowed  for  arbitrarily  hi  the  refining  terms,  avoid- 
ing the  expense  and  delay  of  determining  daily  corrections. 

There  is  no  practical  way  of  ascertaining  the  volatiliza- 
tion loss  except  by  running  " proofs"  made  up  with  known 
quantities  of  silver.  With  high-grade  dore  and  similar 
pure  material  this  is  satisfactory,  but  when  various  im- 
purities are  present,  they  affect  the  volatilization  and  a 
true  proof  would  be  a  very  complicated  affair. 

In  the  case  of  gold  the  fire  assay,  where  the  bullion  is 
scorified  direct  to  a  lead  button,  is  used  as  it  gives  a  higher 
value  than  the  gold  parted  from  the  silver  in  the  combina- 
tion assay,  due  probably  to  filter  losses  and  the  oxidation 
of  some  gold  when  dissolving  the  bullion  in  nitric  acid 
in  the  latter  method.  This  gives  a  return  of  about  99.6 
per  cent  in  the  uncorrected  assay.  The  corrected  gold 
assay  is  very  close  to  full  value.  The  so-called  mercury 
or  sulphuric  acid  method  is  an  equivalent  of  the  fire 
assay. 

b.  Splitting  Limits. — Were  assays  absolutely  accurate 
and  all  laboratories  equally  reliable,  the  question  of 
splitting  limits  would  not  enter  into  a  discussion  of  metal 
losses.  As  such  is  not  the  case  it  is  necessary  to  agree 
upon  some  reasonably  allowable  difference  between  buyer 
and  seller  within  which  assays  shall  be  averaged  or 
" split"  and  beyond  which  an  umpire  assayer  shall  be 
called  upon. 


10  COPPER  REFINING 

In  order  to  guard  against  either  dishonesty  or  care- 
lessness several  simple  precautions  should  be  taken.  In 
the  first  place  assays  should  be  exchanged  simultaneously 
by  mail;  then  all  "reassays"  to  avoid  umpires  should  be 
forbidden;  finally,  splitting  limits  should  be  made  as  close 
as  possible.  Sometimes  the  splitting  limits  are  made  very 
close,  such  as  0.1  per  cent  for  copper,  for  example,  and  then 
only  such  umpires  undertaken  as  may  be  necessary  to 
bring  the  average  results  for  ten  consecutive  lots  within 
splitting  distance. 

The  umpire  should  be  considered  simply  as  a  substitute 
for  one  assayer,  not  as  a  source  of  authority.  When 
two  assays  fall  too  far  apart,  the  assumption  is  that  one 
is  incorrect  and  the  umpire  will  determine  which  one  it  is. 
His  result  should  therefore  be  averaged  with  the  nearest 
original  result  unless  it  is  above  the  higher  or  below  the 
lower  by  more  than  the  splitting  limit,  in  which  case  the 
sample  is  pronounced  uneven  and  the  two  original  assays 
averaged  for  settlement. 

In  many  cases  where  wild  umpires  are  reported  the 
trouble  with  the  sample  can  be  rectified  by  screening  out 
the  fines  in  the  sample  and  then  making  up  a  portion  for 
assay  with  just  the  proper  proportion  of  fines,  the  more 
impure  parts  of  the  alloy  having  been  brittle  and  pul- 
verized under  the  drill. 

It  is  also  well  to  have  two  or  three  competent  umpire 
chemists  agreed  upon,  assigning  umpires,  as  they  become 
necessary,  in  rotation.  This  tends  to  check  any  inac- 
curacies in  the  umpire  laboratory. 

c.  Assay  Errors. — Assays  are  subject  to  three  sources 
of  error:  (1)  those  of  method;  (2)  those  of  procedure; 
(3)  those  of  calculation,  or  clerical  errors.  The  first 
has  already  been  dealt  with  and  may  be  classed  as  a 
voluntary  error;  the  second  may  be  due  to  carelessness, 
inaccurate  apparatus,  such  as  scale  weights  or  burettes, 
or  impure  reagents;  the  third  is  more  serious  than  might 
be  imagined  and  constant  vigilance  is  necessary.  Perhaps 
the  most  common  clerical  error  is  that  of  reporting  double 


METAL  LOSSES  11 

or  half  the  proper  value  in  silver  and  gold  determinations, 
due  to  failure  to  note  the  fraction  of  an  assay  ton  taken  for 
assay. 

D.  SLAGS 

We  have  now  to  pass  from  the  apparent  losses  due  to 
inaccurate  accounting  for  values  in  the  incoming  bullion, 
to  the  actual  metallurgical  losses,  and  the  first  of  these 
is  the  loss  in  the  cupola  slag.  The  anode  and  refining 
furnace  operations  produce  a  certain  amount  of  slag  due  to 
the  reaction  between  the  metallic  oxides  formed  during 
scorification  in  the  furnace  and  silica  and  other  oxides 
present  in  the  furnace  walls  and  hearth,  or  introduced  as 
coal  ashes  blown  over  from  the  fire  box,  or  as  clay  used  in 
luting  up  the  doors,  fettling,  etc. 

In  the  old  days  of  small  furnaces  constructed  entirely 
of  siliceous  material,  from  3  to  4  per  cent  of  the  weight 
of  the  charge  was  made  in  slag.  In  the  large  modern  fur- 
naces constructed  partly  or  wholly  of  basic  or  neutral 
material  and  with  much  better  fuel  economy,  the  slag 
made  is  below  1  per  cent. 

Theoretically  the  slag  should  be  a  very  small  item. 
After  a  charge  is  melted  air  is  blown  in  sufficient  to  oxidize 
part  of  the  copper.  This  cuprous  oxide  in  turn  displaces 
the  impurities  by  oxidizing  them  and  sending  them  into 
the  slag.  The  cuprous  oxide  is  soluble  to  a  certain  extent 
in  the  molten  copper  and  were  no  acids  present  it  would 
be  theoretically  possible  to  skim  off  simply  these  impurities 
with  a  little  mechanically  entangled  copper.  The  actual 
slags  made  run  about  45  per  cent  copper  and  must  be 
treated  in  a  blast  furnace  with  fluxes  such  as  iron  oxide  and 
limestone  to  make  black  copper  unless  a  smelting  plant  is 
connected  with  the  refinery. 

If  some  sulphur  is  introduced  into  this  charge  a  low- 
grade  product  running  about  94  per  cent  copper  may  be 
made  with  a  slag  running  below  1  per  cent  in  copper,  but 
this  gives  a  high  operating  cost  in  the  anode  furnace. 

If  a  high-grade  black  copper  is  made  it  is  very  difficult 


12  COPPER  REFINING 

to  get  a  slag  below  2  per  cent  in  copper.  Taking  a  final 
cupola  slag  of  36  per  cent  silica  and  neglecting  the  small 
quantity  of  silica  introduced  by  the  basic  fluxes,  this  means 
that  every  ton  of  silica  which  is  allowed  to  find  its  way  into 
the  refining  furnace  slags  carries  away  from  56  to  112  Ib. 
of  copper  in  cupola  slag  aside  from  the  stack  loss  in  this 
operation.  It  is  evident  therefore  that  this  great  source 
of  loss  is  worthy  of  the  most  careful  study,  to  the  end 
that  the  least  possible  amount  of  slag  be  made  at  the 
reverberatories. 

E.  STACK  LOSSES 

a.  Anode  Furnace  Stacks. 

b.  Refining  Furnace  Stacks. 

Not  a  great  deal  is  known  about  these  losses.  The  re- 
coveries in  flue  dust  from  waste-heat  boilers  installed  in 
anode  furnaces  show  some  0.07  per  cent  of  the  copper 
treated,  rather  more  of  the  silver  and  less  of  the  gold.  On 
the  other  hand,  bag-house  tests  on  the  gases  escaping  from 
a  high  direct  stack  have  shown  less  than  this  amount. 
The  composition  of  the  bullion  under  treatment  has  doubt- 
less much  to  do  with  these  losses,  as  the  recoveries  on  fur- 
naces treating  cathodes  are  much  less,  due  partly  to  less 
working  of  the  molten  charges  and  partly  to  the  absence  of 
volatile  impurities  which  always  promote  metal  losses. 

c.  Silver  Refinery  Stacks. — The  anode  slimes  consist  of 
the  insoluble  impurities  contained  in  the  anode  and  run  30 
per  cent  to  40  per  cent  in  silver.     The  copper  is  mostly 
leached  out  as  sulphate  and  the  slimes  are  then  melted  and 
subjected  to  a  series  of  oxidizing  operations  until  a  high- 
grade  dore*  is  obtained. 

In  general  about  1  per  cent  of  the  silver  treated  and  about 
0.1  per  cent  of  the  gold  is  recovered  by  various  means  in 
the  flue  system  and  great  progress  has  been  made  in  the 
last  ten  years  in  this  practice. 

Until  the  Cottrell  system  of  electrostatic  precipitation 
was  successfully  applied  to  treating  these  gases,  the  oppor- 
tunities for  serious  undetected  losses  were  very  great  and 


METAL  LOSSES  13 

even  now  there  is  no  point  in  a  metal  loss  investigation 
which  needs,  more  careful  examination. 

The  great  difficulty  is  that  the  actual  losses  made  are 
clearly  discernible  only  after  several  years,  as  the  results 
from  single  yearly  inventories  are  always  more  or  less 
clouded  by  anode  furnace  bottom  absorptions  which  vary 
from  year  to  year,  and  the  punishment  comes  so  long  after 
the  crime  that  it  is  very  easy  to  be  lax  to  the  immediate 
benefit  of  apparent  operating  costs.  Molten  silver  should 
be  handled  like  a  volatile  liquid. 

d.  Cupola  Stack. — This  loss  should  be  relatively  small, 
as  the  operations  are  on  a  small  scale  and  the  charge  in 
reasonably  good  physical  condition.  A  proper  dust  cham- 
ber should  be  installed,  however. 

F.  PROCESS  LOSSES 

a.  Silver  and  Gold  in  Outgoing  Copper. — This  loss  is  a 
perfectly  definite  one  shown  by  daily  assays.     In  general 
the  loss  runs  from  0.5  to  0.8  per  cent  of  the  silver  and  gold 
present  in  the  anodes.     In  the  case  of  low-grade  anodes 
this  is  not  serious,  but  when  rich  material  is  being  treated 
this  loss  becomes  quite  an  item.     The  loss  is  due  to  slimes 
adhering  to  the  cathode  and  is  affected  by  the  current  den- 
sity, the  volume  of  circulation  of  the  electrolyte,  and  the 
degree  of  refining  given  in  the  anode  furnace. 

b.  Silver  in  Outgoing  Gold. — The  Government  does  not 
pay  for  the  silver  in  gold  deposited  unless  the  fineness  is 
low  enough  to  require  refining.     In  the  latter  case  the  re- 
fining charges  imposed  greatly  exceed  the  costs  of  refining 
before  shipment.     The  net  result  is  that  any  silver  con- 
tents are  not  paid  for  and  therefore  constitute  a  metal  loss. 
This  loss  should  not  exceed  two  or  three  parts  per  thousand. 

c.  Gold  in  Outgoing  Silver. — Any  gold  in  outgoing  fine 
silver  is,  of  course,  a  total  loss.     It  should  be  possible  to 
keep  this  loss  down  to  0.1  oz.  per  ton  of  silver. 

d.  Minor  Losses. — These  comprise  values  lost  in  any  by- 
products sold,  such  as  nickel  sulphate,  selenium,  platinum 
etc.,  and  the  value  of  assay  samples  sent  out  without  credit, 


14  COPPER  REFINING 

etc.  These  sources  of  loss  are  relatively  unimportant  but 
should  not  be  overlooked. 

e.  Process  Margins. — These  are  negative  losses  or  gains 
due  to  receiving  more  metal  than  accounted  for,  or  to  ship- 
ping less,  due  to  trade  customs.  In  copper  there  is  no  assay 
margin,  but  there  is  a  gain  of  about  0.07  per  cent  due  to 
the  fact  that  wirebar  copper  runs  but  about  99.93  per  cent 
copper,  while  it  is  credited  as  100  per  cent  against  incoming 
copper  received  on  actual  contents. 

In  the  case  of  silver  the  uncorrected  combination  assay 
allows  the  refiner  about  2.5  per  cent  margin  while  on  ship- 
ments of  fine  silver,  although  the  assay  used  shows  true 
contents,  a  fineness  of  999  is  considered  a  100  per  cent  de- 
livery. With  gold  there  is  a  margin  of  about  0.4  per  cent 
in  the  incoming  bullion  from  the  fire  assay  used,  but  no 
margin  in  the  outgoing  fine  gold. 

G.  HANDLING  LOSSES 

a.  Wind  Losses. — Wind  does  not  cause  as  much  trouble 
in  a  refinery  as  in  a  smelter,  owing  to  the  nature  of  the 
material   handled.     If    the    atmosphere    were    absolutely 
quiet,  however,  it  is  conceivable  that  the  fume  losses  from 
stacks  would  settle  down  within  the  confines  of  the  plant. 
Dust  collected  from  the  roofs  of  the  buildings  is  always  high 
in  grade  although  very  small  in  quantity.     I  have  a  memo- 
randum of  a  sample  of  so-called  dust  swept  up  near  a  Te- 
finery  furnace  building  which  ran  65.81  per  cent  Cu,  49.8 
oz.  per  ton  Ag  and  3.41  oz.  per  ton  Au. 

b.  Theft. — A  refinery  is  always  subject  to  losses  by  pil- 
fering and  a  most  thorough  system  of  patrol,  passes,  etc., 
together  with  a  policy  of  prosecution  of  all  cases  detected 
is  necessary  to  hold  this  in  check.     Copper  is  a  very  easy 
metal  to  sell  as  junk  and  systematic  carrying  away  of  small 
quantities  in  lunch  pails,   etc.,   may  run  into  surprising 
amounts.     In  the  silver  refinery  it  is  customary  to  bond 
the  employee  and  to  have  a  double  set  of  dressing  rooms 
with  different  clothes  for  work  and  for  outside  wear,  the 
men  passing  stripped  before  a  watchman  from  one  room  to 


METAL  LOSSES  15 

the  other  at  change  of  shift.  One  assistance  in  the  detec- 
tion of  silver  and  gold  thefts  lies  in  the  fineness  of  the  prod- 
ucts. When  a  man  offers  999  silver  or  24  carat  gold  for 
sale  he  at  once  arouses  suspicion. 

c.  Solution  Losses. — Although  the  day  has  passed  when 
electrolytes  were  purified  by  running  a  proportion  to  the 
sewer  at  regular  intervals,  there  is  still  plenty  of  oppor- 
tunity for  loss  in  the  handling  of  solutions. 

It  is  very  difficult  to  keep  the  electrolyte  with  its  free 
sulphuric  acid  confined  to  the  circulating  system  on  account 
of  tank  leaks,  overflows,  etc.,  and  it  is  even  more  difficult 
to  build  a  permanent  water-tight  acid-proof  floor  under 
these  tanks. 

Some  of  the  earlier  refineries  suffered  severely  from  such 
losses  and  I  know  of  one  plant  where  it  was  possible  to  dig  a 
well  anywhere  in  the  vicinity  and  pump  out  water  which 
gave  a  profitable  copper  recovery  when  passed  over  scrap 
iron. 

One  good  way  to  keep  track  of  losses  from  this  and  similar 
sources  is  to  keep  a  careful  record  of  sulphuric  acid  move- 
ments in  the  solutions.  Serious  acid  losses  point  at  once  to 
equivalent  copper  solution  losses.  In  plants  where  waste 
liquors  are  worked  up  by  cementation  upon  iron  there  is  an 
additional  opportunity  for  loss  in  undertreated  waste 
liquors. 

d.  Soil  Losses. — Where   a  plant   is  unpaved  there  is 
always  more  or  less  metallic  material  ground  into  the  earth 
and  in  a  smelting  plant  the  top  soil  becomes  in  time  very 
good  ore.     In  a  refinery  there  is  not  the  same  opportunity 
for  loss  owing  to  the  nature  of  the  material  handled  and 
such  a  loss  should  be  negligible  except  in  the  case  of  solution 
losses  mentioned  above  and  that  of  slimes  losses  taken  up 
below. 

e.  Slimes  Losses. — Mechanical  losses  of  slimes   could, 
of  course,  be  assigned  to  the  several  sub-headings  preceding, 
but  the  matter  is  so  important  that  a  separate  paragraph 
has  been  reserved  for  their  discussion. 

Slimes  originate  in  the  tank  house.     They  are  periodi- 


16  COPPER  REFINING 

cally  sluiced  down  into  a  screening  tank  whence  they  are 
usually  transferred  by  pumping  to  a  receiving  tank  in  the 
silver  refinery.  A  certain  amount  is  carried  out  in  the 
cathodes,  a  matter  which  has  already  been  taken  up.  More 
or  less  adheres  to  the  anode  scrap  when  it  is  drawn  and  is 
removed  as  thoroughly  as  possible  by  scrubbing  or  by  high- 
pressure  water-sprays  before  the  scrap  is  sent  to  the  anode 
furnace.  Any  carelessness  here  may  result  in  wind  and  soil 
losses  in  transit  or  up  the  stack  during  melting. 

Tank  leaks  in  the  tank  house  carry  considerable  quantities 
of  slimes  to  the  cellar  floor.  While  most  of  these  values  are 
recovered  by  washing  the  floor,  some  is  absorbed  and  the 
floor  material  has  to  be  smelted  whenever  extensive  repairs 
are  made. 

In  the  silver  building  the  slimes  of  necessity  get  spilled 
around  to  a  certain  extent  and  as  boiling  operations  are 
conducted  in  this  department  the  steam  makes  a  sort  of 
paste  with  them  and  care  must  be  taken  that  all  unneces- 
sary walking  in  and  out  of  this  department  be  avoided, 
and  that  proper  means  for  wiping  shoes  be  provided.  The 
shower  baths  used  by  the  workmen  should  drain  into  the 
general  wash  water  system. 

The  very  best  possible  floor  must  be  provided  in  this 
building.  During  some  changes  in  one  of  the  silver  re- 
fineries some  old  foundations  yielded  up  an  absorption  of 
17,000  oz.  of  silver  and  200  oz.  of  gold. 

The  slimes  before  melting  have  the  appearance  of  black 
mud  and  it  is  very  hard  to  get  workman  to  handle  them  with 
the  care  they  instinctively  give  to  the  silver  sponge  and 
other  metallic  products  produced  later. 

From  the  foregoing  it  will  be  appreciated  that  constant 
thoughtful  attention  in  a  great  many  directions  is  necessary 
to  make  a  minimum  metal  loss.  As  the  precautions  all 
cost  money  it  raises  the  question  whether  the  additional 
saving  due  to  them  pays.  The  answer  to  this  is  that  it  is 
only  by  constant  schooling  in  all  these  precautionary  meas- 
ures that  men  can  be  trained  to  be  really  careful  when  they 
are  on  their  own  responsibility  and  not  under  observation. 


METAL  LOSSES  17 

For  this  reason  the  liberal  use  of  white  paint  and  many  other 
seeming  extravagances  are  justified.  The  same  rigid  care 
and  supervision  in  the  weighing,  sampling,  and  assaying  is 
required  as  hi  the  control  of  slag  and  cathode  losses,  and 
careful  attention  must  be  given  on  general  principles  to  all 
the  apparently  insignificant,  sources  of  indefinite  loss. 

In  general  a  well  conducted  refinery,  operating  on  clean 
bullion  and  accounting  for  contents  as  shown  by  electrolytic 
assay  for  copper,  uncorrected  combination  assay  for  silver 
and  uncorrected  fire  assay  for  gold,  should  show  a  loss  of 
seven  or  eight  pounds  of  copper  per  ton,  at  least  break  even 
on  silver  and  recover  a  small  overage  of  gold. 


CHAPTER  II 


METALS  IN  PROCESS 

Most  electrolytic  refining  processes  have  to  compete  with 
fire  processes  which  are  as  a  rule  less  satisfactory  techni- 
cally, but  which  turn  out  the  product  in  a  very  much 
shorter  time.  This  slow  turn-over  imposes  heavy  interest 
charges  on  account  of  the  large  investment  in  plant  which 
it  implies  and  from  the  necessity  of  financing  the  metals  in 
process.  The  plant  investment  and  the  amount  of  metals 
locked  up  are  closely  related  to  the  current  density  em- 
ployed, but  as  the  power  required  increases  nearly  as  the 
square  of  the  current  density,  a  point  is  soon  reached  where 
the  cure  is  worse  than  the  disease. 

TABLE  1 


Metal 

Normal  Pre-  Wai- 
Market   Value 
Dollars  per  Ib. 

Electrochemical 
Equivalent 
Grams  per 
Ampere-hour 

Quotient 
X104 

Iron 

0  01 

1.04 

1 

Lead  
Zinc 

0.04 
0.06 

3.86 
1.22 

1 

5 

Copper 

0  15 

1.19 

13 

Nickel.  
Silver                    

0.30 
9.00 

1.09 
4.03 

29 
232 

Gold 

300  00 

2.45 

12,700 

The  case  varies  greatly  with  the  metal  under  considera- 
tion, depending  upon  its  electrochemical  equivalent  and 
market  value..  Some  idea  of  this  variation  can  be  obtained 
from  an  inspection  of  Table  1,  which  takes  the  ratio  of 
these  two  factors  as  a  measure  of  the  relative  costs.  Al- 
though this  table  is  inexact  in  that  it  ignores  voltage  and 
other  minor  factors  it  shows  how  in  the  case  of  lead  a  handi- 
cap has  been  removed  enabling  electrolysis  to  compete 

18 


METALS  IN  PROCESS  19 

with  the  best  of  fire  processes,  while  gold  is  electrolyzed 
in  but  few  plants  not  under  Government  auspices,  where 
locked  values  are  not  of  any  importance. 

The  question  is  somewhat  complicated  by  the  fact  that 
the  bullion  being  electrolyzed  is  generally  a  mixture  of  two 
or  more  metals,  each  of  which  is  to  be  recovered.  The 
refining  of  blister  copper  carrying  silver  and  gold  will  serve 
as  an  example  upon  which  to  base  discussion  of  the  general 
problem. 

Copper  refining  contracts  specify  certain  permissible 
elapsed  times  for  the  return  of  or  payment  for  the  values 
in  the  bullion  to  be  refined,  varying  from  45  to  75  days  for 
the  copper  and  from  60  to  90  days  for  the  silver  and  gold. 
It  is  evident  that  the  earlier  the  metals  can  be  put  on  the 
market  the  sooner  their  cash  value  can  be  put  to  earning 
money  elsewhere  and,  therefore,  that  each  day  that  the 
process  locks  up  the  values  has  a  definite  cash  value.  This 
is  not  greater  than  6  per  cent,  however,  as  the  banks  con- 
sider warehouse  certificates  based  on  metals  in  process 
satisfactory  collateral  for  loans  up  to  a  high  percentage  of 
their  market  value  in  ordinary  times.  Taking  copper  at 
fifteen  cents  a  pound,  the  interest  value  of  a  sixty-day  re- 
fining allowance  is  three  dollars  a  ton,  and  each  ton-day 
costs  five  cents,  with  corresponding  values  for  the  silver  and 
gold  contents.  The  cost  of  refining  is  made  up  of  operating 
expenses,  metal  losses  and  metal  interest,  and  it  is  the 
purpose  of  this  chapter  to  examine  the  last. 

Metals  are  tied  up  in  process  for  three  quite  different 
reasons  which  might  be  stated  as  commercial,  technical, 
and  balancing.  Commercial  policy  makes  it  desirable  to 
hold  incoming  bullion  at  times  before  putting  it  into  process 
until  certain  conditions  are  fulfilled,  such  as  agreement  of 
assays,  or  such  a  state  of  weather  as  shall  obviate  the  neces- 
sity of  moisture  allowance  when  weighing;  or  fluctuations  in 
the  metal  market  may  make  it  advantageous  to  go  to  extra 
operating  costs  to  cut  down  what  normally  would  be  con- 
sidered proper  stocks  in  process  or  to  delay  the  purchase 
of  metals  lost  in  process;  finally  it  is  good  policy  to  endeavor 


20  COPPER  REFINING 

to  carry  reasonable  stocks  of  refined  shapes  on  hand  so  as  to  be 
able  to  adjust  shipments  to  unexpected  changes  in  orders, 
steamer  sailings  and  the  like.  It  is  quite  reasonable  to 
have  10  per  cent  of  the  total  tie-up  in  copper  due  to  such 
causes. 

Then  we  have  the  values  locked  up  for  technical  reasons 
which  may  be  classified  as  metals  truly  in  active  process, 
metals  circulating  in  slags,  etc.,  metals  permanently  tied 
up  in  furnace  bottoms,  etc.,  and  metals  forever  lost  to  be 
charged  into  operating  costs  as  metal  losses. 

Finally  there  are  the  stocks  of  pig,  anodes  and  cathodes 
which  it  is  necesary  to  carry  in  order  to  insure  smooth, 
continuous  operation  regardless  of  minor  irregularities  in 
arrival  of  pig  or  breakdowns  in  some^  parts  of  the  process. 
To  these  the  name  of  " balancing"  metals  has  been  given 
for  want  of  a  better  term. 

The  total  metals  tied  up  are,  of  course,  shown  by  the 
balances  due  customers  on  the  metal  books  of  the  plant, 
but  these  figures  give  no  data  of  value  in  the  analysis  or 
control  of  the  values.  It  is  also  customary  to  take  periodi- 
cal physical  inventories  of  metals  on  hand  in  order  to 
determine  accurately  the  metal  losses  for  the  preceding 
term.  This  indicates  the  metal  tied  up  in  each  part  of  the 
plant  but  unfortunately  in  order  to  take  a  proper  inventory 
without  shutting  down  the  process  it  is  necessary  to  intro- 
duce temporarily  radical  changes  in  operation  which  greatly 
increase  the  tie-up.  It  is,  therefore,  necessary  to  rely 
upon  a  thorough  understanding  of  the  sources  of  locked 
values  and  of  their  relative  importance  and  a  constant 
scrutiny  of  conditions  in  order  to  control  this  important 
factor  in  the  total  cost  of  operating  a  copper  refinery. 

Classification. — Figure  3  gives  a  systematic  classification 
of  the  sources  of  locked  values  which  we  shall  proceed  to 
take  up  item  by  item.  The  items  under  the  heading  "  Com- 
mercial" may  be  dismissed  with  the  arbitrary  assumption 
of  five  days  delay  of  copper,  four  days  of  silver  and  three 
days  gold. 


METALS  IN  PROCESS 


21 


Metals  in  Process. — Turning  next  to  the  delays  for 
technical  reasons  and  taking  first  " metals  in  process," 
Fig.  4  gives  the  direct  course  of  the  metals  through  the 
process  ignoring  all  diversions. 


lyocked  Values 


Commerical 


lee 


Pig 
Embargoes 


Wirebar 
Stocks 


Miscellaneous 


meal 


Balancing 


Slags       Pig     Anodes  Cathodes 

Slimes  Starting  Sheets 


Metals  Circulating 


Metals  i 


Molds 


Slags 


Flue       Liquors 
Products  to 

Anode 


Process 


Metals 
Withdrawn 


Cathode    Refining 

to  Anode 

Wirebar     Slimes 


Anode 

to 
Cathode 


Metals 
Lost 


Chapter  1 


Liquors      Furnace 
Dottoms 


Moulds 
Etc 


Bus-bars 
Etc 


FIG.  3. — Analysis  of  sources  of  locked  values. 

The  weighing,  sampling,  and  handling  of  pig  to  the  anode 
furnace  properly  occupies  one  day.  The  refining  and  cast- 
ing into  anodes  is  a  twenty-four  hour  operation.  In  the 
tank  house  conditions  vary  greatly  in  different  plants,  but 
representative  eastern  practice  may  be  taken  as  14  days 
cathodes  and  28  days  anodes,  making  an  average  age  of  21 
days  for  the  two  sets  of  electrodes.  The  refining  furnace  is 
operated  on  the  same  basis  as  the  anode  furnace,  calling  for 
one  day.  Certain  parts  of  the  plant  are  shut  down  on 
Sundays,  but  this  is  assumed  to  have  no  effect  upon  the  final 
date  of  delivery  of  refined  copper. 


22 


COPPER  REFINING 


In  the  silver  refinery  the  length  of  time  required  to 
recover  dore  from  the  slimes  will  depend  greatly  upon  the 
impurities  present,  but  is  not  likely  to  be  less  than  three 


Slimes 


Market  Shapes 
Copper 


Silver 


Gold 
FIG.  4. — Direct  course  of  metals  through  plant. 

days  for  dissolving  the  copper  in  the  slimes  and  six  days 
for  furnacing  them. 

In  the  parting  plant,   assuming  electrolytic  parting  is 
used,  the  anodes  will  probably  last  three  days  and  the 


METALS  IN  PROCESS  23 

cathode  crystals  be  cleaned  up  every  few  hours,  making  an 
average  of  one  and  a  half  days.  The  washing  and  melting 
of  the  silver  sponge  takes  but  a  few  hours,  say  half  a  day. 
The  gold  mud  would  be  cleaned  every  three  days  and  the 
boiling  and  melting  take  two  days. 

We  can  now  sum  up  the  time  required  as  shown  in 
Table  2. 

TABLE  2 

Copper    Silver  Gold 

Weighing  and  sampling 1  1  1 

Anode  furnaces 1  1  1 

Tank  house 21  21  21 

Refining  furnaces 1 

Boiling  slimes 3  3 

Furnacing  slimes 6  6 

Parting  dore 1.5  3 

Melting  silver 0.5 

Refining  gold 1.5 

Melting  gold 0.5 

Total  days 24  34  37 

These  figures  show  first  that  but  about  half  the  time 
allowed  for  refining  is  required  by  a  straight  passage  through 
the  process,  thereby  giving  the  goal  we  should  strive  to 
approach,  and  second  that  as  would  be  expected  the  slow 
item  is  the  electrolysis.  This  time  can  be  lessened  by  cut- 
ting down  the  age  of  the  anodes  or  of  the  cathodes  or  of  both. 
Too  light  an  anode  increases  handling  costs  and  the  per- 
centage of  anode  scrap  made,  while  too  light  a  cathode  runs 
up  disproportionate  charges  for  making  starting  sheets  and 
decreases  the  productive  hours  of  the  tanks,  which  have  to 
be  cut  out  while  drawing  copper.  As  before  stated  the 
power  charges  mount  very  rapidly  with  increasing  current 
density. 

Metals  Circulating. — The  sample  diagram  shown  in  Fig. 
4  would  become  an  almost  undecipherable  confusion  of 
lines  were  it  expanded  to  indicate  every  movement  of 
circulating  metals.  As  we  proceed  from  pig  to  cathode  a 
fraction  is  diverted  at  each  stage  of  the  process  and  sent  back 
to  an  earlier  one. 


24  COPPER  REFINING 

Each  of  the  furnace  processes  may  be  illustrated  by  the 
first  diagram  in  Fig.  5  and  both  the  copper  electrolysis  and 
the  dore  parting  by  the  second  except  that  there  are  no 
starting  sheets  in  electrolytic  parting  processes,  the  silver 
being  deposited  in  non-adherent  crystals  upon  a  fixed 
carbon  or  silver  cathode. 


Metal  Los«  I  \  LiquorB 


Metal  Loss 


Bottom  v  ' 

Absorption,  etc 


FIG.  5. — Typical  metal  movements. 

The  return  arrows  indicate  circulating  metals  and  the 
straight  ones  metals  locked  up  continuously.  The  two 
cases  are  alike  in  their  effect  upon  the  process  except  that 
the  former  varies  directly  with  the  output  while  the  latter 
are  more  or  less  constant  for  a  given  plant  regardless  of 
output.  (Metals  actually  lost  are  a  third  case,  but  the 
quantities  involved  are  small  and  are  generally  adjusted 
by  systematic  purchases  so  as  to  have  but  a  negligible 
effect  upon  the  time  required  for  refining  and  they  will  be 
ignored.)  Liquors  partake  of  both  characteristics,  as  a 
certain  stock  is  required  for  starting  up  the  plant  while 
another  quantity  circulates  through  the  purifying  depart- 
ments, etc. 

Returning  to  the  plan  laid  down  in  Fig.  3  we  shall  con- 
sider the  metals  circulating  in  slags,  flue-products  and 
liquors. 

Anode  Furnaces. — The  first  and  simplest  case  is  that 
presented  by  the  anode  furnaces;  first  because,  if  we  ignore 
the  small  amount  of  values  circulating  in  the  drillings 
from  sampling,  the  anode  furnaces  divert  the  first  fraction 
from  the  entering  pig,  and  simplest,  because  the  slag  and 
flue  products  made  return  directly  to  the  anode  furnaces 


METALS  IN  PROCESS  25 

themselves  after  being  reduced  to  pig  in  a  cupola,  giving  a 
single  fraction  to  consider. 

The  quantity  of  slag  made  by  an  anode  furnace  varies 
with  the  purity  of  the  pig  from  less  than  1  per  cent  of  the 
charge  to  several  per  cent.  The  slag  itself  will  run  from 
35  to  45  per  cent  copper  and  the  silver  and  gold  slagged 
will  be  proportionately  less  than  the  copper.  We  shall  as- 
sume that  we  have  made  1.5  per  cent  of  the  copper  contents 
•of  charge  in  slag  running  40  per  cent  copper  and  that  the 
silver  and  gold  slagging  ratio  is  0.85  and  0.45  respectively, 
with  copper  at  1.00.  In  flue  dust  we  shall  recover  about 
0.07  per  cent  of  the  copper  and  say  0.1  per  cent  of  the  silver 
and  0.05  per  cent  of  the  gold  treated.  Then  we  have  other 
flue  products  such  as  the  cobbing  from  repairs,  and  the 
quantity  of  this  also  varies  greatly  with  the  character  of 
the  bullion  under  treatment.  We  shall  assume  it  to  amount 
to  25  per  cent  of  the  metals  in  the  slag.  Then  there  is  a 
negligible  amount  of  bad  production,  scale,  etc.,  charged 
back  next  day.  We  have  therefore  as  a  measure  of  the 
apparent  shrinkage  of  the  values  in  process,  ignoring  actual 
metal  losses,  the  figures  given  in  Table  3. 

TABLE  3. — ANODE  FURNACE  DIVERSIONS 

Copper  Silver       Gold 

Charged  to  furnace. ...  1 .0000  1 .0000  1 .0000 

Slag  contents 0.0060  0.0051  0.0027 

Flue  dust ...   0.0007  0.0010  0.0005 

Cobbing,  etc 0.0015  0.0013  0.0007 


Total  diversions 0.0082     0.0074     0.0039 

Net  output .  .  .   0 . 9918     0 . 9926     0 . 9961 

The  actual  retreatment  of  these  by-products  does  not 
take  over  a  day,  the  process  being  simply  reduction  to  pig 
copper  in  a  small  cupola.  In  practice  a  great  deal  of  time 
is  lost  in  storage  before  retreatment,  but  this  will  be  taken 
up  later  under  "  balancing "  the  process. 

Tank  House. — The  tank  house  has  three  circulating 
items;  airode  scrap,  starting  sheets,  and  liquors. 


26  COPPER  REFINING 

Anode  scrap  is  the  part  of  an  anode  remaining  after  it 
has  been  eaten  away  by  the  current  until  but  a  skeleton 
remains  hanging  from  the  supporting  lugs.  These  lugs 
generally  account  for  from  3  to  5  per  cent  of  the  total  weight 
of  the  anode  and  this  percentage  represents  the  lowest  pos- 
sible anode  scrap  unless  lugless  anodes  are  used  as  shown  in 
Fig.  13. 

In  practice  the  scrap  made  varies  between  7  and  20  per 
cent.  A  poorly  refined  anode  will  dissolve  unevenly  and 
cause  heavy  scrap,  while  careful  filling  of  the  "holes" 
caused  by  drawing  individual  anodes  which  work  out  first 
will  lessen  the  final  scrap  made.  The  best  policy,  however, 
is  to  refine  the  anodes  thoroughly  and  take  no  chances  on 
disturbed  current  distribution  in  the  tanks  due  to  "lace" 
anodes,  so  that  a  normal  percentage  of  scrap  will  be  about 
14  per  cent  of  the  weight  of  the  anodes  charged.  This 
scrap  is  subject  to  a  charge  of  28  days  in  the  tank  house 
plus  the  time  required  for  the  anode  cycle. 

The  starting  sheets  are  cathodes  which  are  deposited 
upon  greased  copper  plates  and  stripped  as  a  thin  sheet 
when  they  are  about  twenty-four  hours  old.  They  are  used 
as  mother  sheets  upon  which  to  build  up  the  cathodes 
proper.  As  the  cathodes  are  drawn  in  fourteen  days,  one 
twenty-eighth  of  the  number  of  tanks  (each  copper  plate  or 
"starting  blank"  yields  two  starting  sheets)  must  be  turned 
over  to  this  service.  Actually  a  little  more  than  this  is 
required  for  the  manufacture  of  loops  by  which  to  hang  the 
cathodes  and  to  allow  for  bad  sheets  and  sometimes  for 
wider  spacing.  The  production  of  bad  sheets  will  be  con- 
sidered negligible  and  one  twenty-fifth,  or  4  per  cent,  of 
the  output  will  be  considered  to  be  in  starting  sheets  which 
have  been  set  back  two  days,  one  day  depositing  and  one 
for  attaching  loops,  etc.,  by  this  diversion. 

The  copper  tied  up  in  the  electrolyte  comes  chiefly  under 
the  head  of  "metals  withdrawn,"  but  a  certain  amount  cir- 
culates between  the  tank  house  and  the  silver  building  and 
the  tank  house  and  the  purifying  department.  The  first 
of  these  is  a  small  item.  The  slimes  are  pumped  over  to 


METALS  IN  PROCESS  27 

the  silver  building  as  a  soup  carrying  perhaps  5  per  cent 
solids,  but  as  the  slimes  amount  to  but  about  one  and  a 
quarter  per  cent  of  the  anodes  and  as  the  electrolyte  carries 
but  3  per  cent  copper,  the  total  copper  involved  in  the  move- 
ment is  but  three  quarters  of  a  per  cent  of  the  anode  weight. 
The  slimes  are  settled  out  and  the  liquor  returned  in  about 
twenty-four  hours. 

The  proportion  of  liquor  required  for  purifying  depends, 
of  course,  upon  the  impurities  present  in  the  anode.  In 
the  first  place  there  is  a  certain  amount  of  copper  dissolved 
chemically  from  the  electrodes  which  would  cause  an 
accumulation  of  copper  in  the  electrolyte  were  there  no 
compensating  factors.  If  the  anodes  are  very  free  from 
impurities,  insoluble  anode  tanks  must  be  operated  to  re- 
cover about  2  per  cent  of  the  cathode  output  in  order  to 
strike  a  balance. 

On  the  other  hand  if  nickel  is  present  in  the  anodes  this 
dissolves  electrochemically  but  is  not  deposited  at  the  cath- 
ode, so  that  if  much  is  met  with  a  condition  of  actual 
insufficiency  of  copper  in  the  electrolyte  may  obtain,  calling 
for  the  use  of  shot  towers. 

Antimony  may  be  precipitated  chemically  as  oxychloride 
and  arsenic  electrolytically  in  the  insoluble  anode  tanks. 
Iron  should  be  adequately  removed  in  the  anode  furnaces. 
Nickel,  however,  accumulates  and  must  be  removed  by 
evaporating  liquors. 

It  will  thus  be  seen  that  quite  different  conditions  may 
exist  at  different  plants  or  at  the  same  plant  at  different 
times.  For  our  case  we  may  assume  that  the  equivalent  of 
1  per  cent  of  the  copper  in  the  anode  has  to  be  regularly 
removed.  Practically  all  of  the  copper  contents  of  this 
solution  will  be  eliminated  in  the  three  successive  sets  of 
insoluble  anode  tanks  through  which  it  will  be  passed  as  a 
first  stage  in  the  purification  and  we  need  not  follow  this 
further.  We  may  assume  that  these  tanks  will  be  cleaned 
once  in  ten  days  and  the  product  sent  to  the  cupola. 

Then  we  have  the  copper  in  the  anode  slimes  returned 
from  the  silver  building.  This  amounts  to  about  0.20  per 


28  COPPER  REFINING 

cent  of  the  copper  in  process,  or  say  0.17  per  cent  of  the 
anode  weight.  It  is  returned  to  the  tank  house  as  copper 
sulphate  and  has  about  a  ten-day  cycle. 

Summarizing,  we  have  for  a  unit  of  anode  copper  enter- 
ing the  tank  house  the  diversions  given  in  Table  4. 


TABLE  4. — DIVERSIONS  FROM  TANK  HOUSE 

Copper  Route 

Entering 1 . 0000 

Anode  scrap 0 . 1400  Anode  furnaces 

Starting  sheets 0 . 0400  Tank  house 

Copper  in  slimes 0.0017  Tank  house 

Electrolyte  in  silver  building 0 . 0075  Tank  house 

Electrolyte  to  purifiers O.OlOO  Cupola 


Total  diversions 0. 1992 

Net  output .   0.8008 


Refining  Furnaces. — The  refining  furnaces  are  similar  to 
the  anode  furnaces  as  regards  circulating  metals  except  in 
that  the  quantity  of  by-products  made  is  smaller  owing  to 
the  practical  absence  of  impurities  and  that  a  handicap  ex- 
ists owing  to  the  necessity  of  sending  these  back  through 
the  cupola  to  the  very  beginning  of  the  process.  The  values 
assumed  for  the  anode  furnace  may  be  scaled  down  as  given 
in  Table  5. 


TABLE  5. — DIVERSIONS  FROM  REFINING  FURNACE 

Copper  Route 

Charged  to  furnace 1 . 0000 

Slag 0.0030  Cupola 

Flue  dust 0 . 0001  Cupola 

Cobbing .... ... ...;... 0.0010  Cupola 

Bad  production 0.0100  Refining  furnace 

Molds 0.0200  Refining  furnace 


Total  diversions 0.0341 

Net  production 0 . 9659 


METALS  IN  PROCESS 


29 


The  bad  production  consists  of  defective  castings  which 
are  charged  back  into  the  furnace  next  day.  The  molds 
may  be  considered  also  as  returned  next  day,  the  balance  of 
the  period  by  their  use  being  considered  under  "  metals 
withdrawn."  We  are  assuming  that  the  anode  furnaces 
use  cast  iron  and  not  refined  copper  molds.  The  small 
amounts  of  silver  in  gold  in  the  refining  furnace  by-products 
are  ignored. 

Silver  Refinery. — We  come  next  to  the  silver  refinery. 
The  anode  slimes  carry  all  the  silver  and  gold,  neglecting 
losses,  and  about  0.2  per  cent  of  the  copper,  as  already 
stated.  For  our  purposes  the  small  amount  of  copper 
entering  the  dore  furnaces  may  be  neglected. 

The  next  step  in  the  treatment  of  the  slimes  is  a  rever- 
beratory  smelting.  The  slimes  are  melted  and  tend  to 
form  three  distinct  layers,  bullion,  matte,  and  slag.  The 
slag  is  skimmed  off  and  the  matte,  if  small  in  quantity, 
as  it  should  be  if  the  copper  has  been  properly  removed  in 
the  slimes  boiling  tanks,  broken  up  by  blowing.  The  foul 
bullion  is  then  refined  to  high-grade  dore  by  cupeling  with- 
out the  addition  of  lead.  The  final  by-products  are  two 
slags,  flue  dust  and  cobbing,  all  of  which,  with  the  possible 
exception  of  the  cobbing  which  may  go  to  the  cupola,  after 
sorting  out  metallics,  are  sent  to  the  anode  furnace.  We 
shall  assume  the  data  stated  in  Table  6. 


TABLE  6. — DORE  FURNACE  BY-PRODUCTS 


' 

Weight 

Assay  —  02 

.  per  ton 

Silver 

Gold 

Slimes 

2000 

12000 

100 

Poor  slag  
Rich  slag  

300 
500 

800 
1,200 

1 
6 

Flue  dust 

200 

1  200 

1 

Cobbing  

40 

1,200 

6 

30 


COPPER  REFINING 


The  dore  is  sent  to  the  electrolytic  parting  plant,  but 
as  there  is  neither  anode  scrap,  nor  starting  sheets,  assum- 
ing the  Thum-Balbach  system  with  horizontal  electrodes 
is  employed,  the  only  circulating  values  are  in  the  wash 
waters  from  the  crystal  silver  and  these  are  so  small  that 
they  may  be  neglected  here.  The  same  is  true  of  the  slags 
from  the  crucible  melting  of  the  gold  and  silver. 

There  is,  however,  some  circulating  silver  from  the 
refining  of  the  gold  anode  mud.  This  mud  may  be  con- 
sidered half  gold  and  half  silver,  so  a  quantity  of  silver 
just  equal  to  the  gold  output  is  dissolved  in  the  boiling 
kettles  and  sent  back  to  the  slimes  boiling  tanks.  As  we 
have  assumed  slimes  with  12,000  oz.  per  ton  of  silver  and 
100  oz.  per  ton  of  gold  this  represents  a  diversion  of  0.0083 
or  a  net  output  of  0.9917. 

We  can  now  build  up  a  schedule  for  the  reverberatory 
dore  furnace.  This  is  given  in  Table  7. 


TABLE  7. — DORE  FURNACE  DIVERSIONS 


Silver 

Charged  to  furnace 1 . 0000 

Rich  slag 0.0250 

Poor  slag. . . ' 0.0100 

Flue  dust 0.0100 

Cobbing 0.0020 


Gold 
1.0000 
0.0150 
0.0015 
0.0010 
0.0012 


Route 

Dore  furnace 
Anode  furnace 
Dore*  furnace 
Anode  furnace 


Total  diversions 0.0470    0.0187 

Net  output 0.9530    0.9813 


Summary  of  Circulating  Metals. — We  are  now  in  a 
position  to  revise  Table  2  by  the  application  of  the  diver- 
sion factors  for-  circulating  metals  in  order  to  show  what 
proportion  of  the  metals  entering  the  process  emerge 
directly  in  the  period  given  in  that  table.  This  summary 
of  diversions  is  given  in  Table  8. 

The  figures  of  Table  8  show  that  but  77  per  cent  of 
the  copper,  81  per  cent  of  the  silver,  and  84  per  cent  of  the 
gold  entering  the  plant  go  directly  through  the  process  in 


METALS  IN  PROCESS 
TABLE  8. — SUMMARY  OF  DIVERSIONS 


31 


Copper 

Silver 

Gold 

Factor 

Cumu- 
lative 

Factor 

Cumu- 
lative 

Factor 

Cumu- 
lative 

Weighing  and  sampling 
Anode  furnaces  
Tank  house 

1.0000 
0.9918 
0.8008 
0.9659 

1.0000 
0.9918 
0.7942 
0.7671 

1.0000 
0.9926 
0.8600 

0.9530 
1.0000 
1.0000 
0.9917 
1.0000 

1.0000 
0.9926 
0.8556 

0.8154 
0.8154 
0.8154 
0.8086 
0.8086 

1.0000 
0.9961 
0.8600 

0.9813 
1.0000 
1.0000 
1.0000 
1.0000 

1.0000 
0.9961 
0.8566 

0.8406 
0.8406 
0.8406 
0.8406 
0.8406 

Refining  furnaces  
Furnacing  slimes  

Parting  Dore 

Melting  silver  

Refining  gold  ,  , 

Melting  gold 

Net  product  

0.7671 

0.8086 

0.8406 

the  24,  34,  and  37  days  shown  by  Table  2,  the  balance 
being  diverted  into  by-products  which  require  retreat- 
ment.  The  result  of  this  is  that  the  actual  metals  in 
process  of  treatment  at  various  stages  is1  greater  than  the 
pig  receipts  or  cathode  output  by  the  amount  of  these 
circulating  metals.  It  is  also  obvious  that  as  these  re- 
treated metals  start  through  the  process  a  second  time  a 
portion  is  again  diverted.  There  are  several  ways  of 
evaluating  this,  the  simplest  of  which  is  probably  to  work 
out  several  terms  of  a  continued  fraction,  thereby  ascer- 
taining the  total  metals  passing  through  each  step  of  the 
process. 

Balancing  Metals. — In  order  to  get  the  time  element  we 
must  consider  the  " balancing  metals,"  or  stocks  held  at 
various  points  hi  the  process  hi  order  to  insure  continuity 
of  operation  without  disproportionate  cost.  This  is  done 
graphically  in  Fig.  6,  where  the  numbers  represent  days. 
Where  they  occur  between  the  rectangles  representing 
steps  in  the  process  they  stand  for  the  average  number  of 


32 


COPPER  REFINING 


days  output  at  that  point  held  in  stock;  where  they  occur 
inside  or  alongside  the  rectangles  they  represent  the  days 
in  process  shown  by  Table  2. 

We  can  now  proceed  to  construct  the  diagrams  show- 
ing the  unit-days  lost  at  each  step  and  this  has  been  done 


10 


Silver 


FIG.  6. — Days  required  for  each  step  including  balancing. 

separately  for  the  copper,  silver  and  gold  in  Figs.  7,  8,  and 
9.  Each  vertical  row  in  these  figures  represents  one  term 
of  the  continued  fraction.  Three  terms  are  sufficient  to 
approximate  the  total.  In  order  to  simplify  the  diagram 
the  cupola  has  been  omitted,  although  allowed  for  in  the 
days  given,  and  some  items  have  been  averaged  in  groups. 


METALS  IN  PROCESS 


33 


The  figures  are  units  of  weight  times  days  equals  unit- 
days,  and  as  one  unit  of  weight  is  shown  entering  the  process 
the  sum  of  the  unit-days  is  the  actual  total  days  that  unit 
is  in  process. 

Metals  Withdrawn. — We  have  finally  to  consider  the 
" metals  withdrawn"  by  liquors,  furnace  bottoms,  con- 
ductors, and  molds.  These  are  easily  translated  into 


Weighing 
and 
Sampling 

Anode 
Furnaces 

1.00  Copper 
1 

l.OOxl.  CO  -1.00 

J 

1.00x2.00-2.00 
0.006x21.0=0.17                               nnmi  ~i  « 

l.OOxl. 

00-1.00                               O.lSxl-UO-U.18             o-03       J0.026xl.00-0.08 

Tank 
House 

Refining 
Furnaces 

0.99x1.00-0.99  ^  T            0.16x1.00-0.16    /^  f 

/rvV          /fr'     1                                   /fc^          /-*? 

0.902x21.0-20.83   *         /#        0-159x21.0-3.34                /&      0-0257x21.0-0.54 
^         /$•        1                            \.<7.s,    /&        \                          1 

^^V  '                          0-127x2.  00  -S(^/a, 
0.79x2.00-  l^\                            025         /V«!fl                     0.021x2.00-0.04 

158  7V     L.  /  N?      I 

0.70x1. 

00-0.79                         ^    0.20x1.00  =  0.20       Negligible    ]  0.034x1.  00-  n.na 

n  no.  -i   rfi     n  (TO                                                                        "H 

1         "    1             1 

0-77                    -h                      0-20                    -J-                      0.03       =    1.00 

1.00        Entering  Plant                                          24-00       Days  Direct 
1.18        Entering  Anode  Furnaces                        3.93       Days  Circulating 
1.18              ••         Tank  House                               6.67       Days  Balancing 
1.02             ..         Refining  Furnaces                  34.60       Days  Total 

FIG.  7. — Days  in  process — copper. 

unit-days  by  considering  how  long  the  plant  would  have  to 
run  in  order  to  fill  these  storage  vaults. 

Liquors. — We  have  two  main  bodies  of  liquor  to  con- 
sider, the  copper  sulphate  electrolyte  in  the  tank  house  and 
the  silver  nitrate  electrolyte  in  the  parting  plant,  the  latter 
carrying  negligible  amounts  of  copper,  however. 

In  the  tank  house  in  addition  to  the  liquor  actually 
in  the  tanks  there  is  a  large  additional  quantity  in  the 
piping  system  and  sumps.  In  general  in  a  large  plant 


34 


COPPER  REFINING 


there  is  required  about  20  Ib.  of  liquor  per  pound  per  day 
of  copper  treated.  As  the  electrolyte  runs  about  3  per 
cent  in  copper  this  is  equivalent  to  saying  that  0.6  days 
output  of  the  plant  will  be  required  to  stock  up  the 
electrolyte. 


Weighing 
and 
Sampling 

Anode 
Furnaces 

Tank 
House 

Boiling 
Tanks 

Dore 
Furnace 

Parting 
Plant 

Crucible 

1.00  Silver 

00-2.00 
0.0074x21.0 

0  001i21  00          •  •• 

1.00x1.00=1.00 

1.00x2. 

l.OOxl.00-1.00 

0.16x1.00-0.16          _o.02     >    0.025x1.00-0-03 

-  0.16     J 

0.99x1.00=0.99  /§>l 

0.16x1.00-0.16    /&/ 
/ff  I                0.025x1.00.0.03 

/&  1       \ 

0.993x21.0-20.85 

)0-0.85/^ 

r 

w 

0.155x21.0-3.26  K^      /^    0.025x21.0-0-53 

00-0.02 

0.85s  l.< 

/<=> 

0.18x1.00-0.13^ 

0.022x1. 

0.856x3.00-2.57 

0.14x3.60=0.42           1^      ^  0.023x3.00-0.07 

0.856x4 

0.14x4.  10  =0.5CX^^ 

0.023x4.00-0.09 

0.856x6.00=5.12 

Ab.OSOxS^ob  0.17x6.00=1.02   K^-oos*840.029xC.OO-0.17 

00-0.12 
-0 

=  1.00 

0.807x4.00  =  3.23 

0.16x4.00=0.64 

0.029x4. 

0.809x1.50-1.21 

0.16x1.50-0.24                           0.029x1-50-0.04 

0.809x0-0 

0.16x0-0 

0.029x0 

0.809x0.50-0.40 

0.16x0.  50=-  0.08                           0.029x0.50  =0.01 

0.81                    4- 

016                     +                      0.03 

1.00      Entering  Plant                                        34.00      Days  Direct 
1.18             ••         Anode  Furnaces                      4.21         ..     Circulating 
1.18             ,,         Tank  House                            14-01         .,     Balancing 
'1.06            „         Dore'  Furnace                       52.22         .,     To^ 

FIG.  8. — Days  in  process — silver. 

In  the  parting  plant  this  factor  is  a  little  less,  say  0.4 
day,  as  there  is  not  the  circulating  system  to  allow  for. 

Furnace  Bottoms. — There  are  four  main  furnace  bottom 
absorptions,  anode,  refining,  cupola,  and  dore.  The  anode 


METALS  IN  PROCESS 


35 


and  refining  furnaces  are  alike  except  that  some  18  per 
cent  more  capacity  is  required  in  the  anode  department  as 
shown  in  Fig.  7,  and  that  the  refining  furnace  ties  up  no 


Crucible 


O.S4 


=      1.00 


1.00  Entering  Plant 
1.13  •>         Anode  Furnaces 

1.16  "          Tank  House 

1.02  ••         Core  Furnaces 


37-00  Day*  Direct 
3-66          ..     Circulating 
13.54         .,      Balancing 
54.20          ,,      Total 


FIG.  9. — Days  in  process — gold. 

silver  and  gold.  An  anode  furnace  will  absorb  from  100 
to  400  Ib.  of  copper  per  square  foot  of  hearth  area,  depend- 
ing upon  the  age,  thickness  and  composition  of  the  bottom. 


36  COPPER  REFINING 

It  is  customary  to  start  the  soaking  of  a  new  bottom 
with  material  lean  in  silver  and  gold  and  in  general  the 
relative  absorption  of  silver  is  lower  than  the  amount  to 
be  expected  from  the  grade  of  bullion  treated;  this  ratio 
changes  with  age,  however,  and  gold  in  particular  seems 
to  concentrate  in  the  bottom. 

For  our  purpose  we  may  assume  0.40  days  copper  for  the 
complete  absorption  of  the  furnace  and  flues  of  a  refining 
unit.  This  will  become  0.47  days  for  an  anode  furnace. 
The  silver  will  be  say  0.25  and  gold  0.40.  These  figures  must 
be  increased  about  30  per  cent,  however  for  reserve  capacity. 

The  cupola  bottom  absorption  is  relatively  negligible. 
This  is  due  to  the  fact  that  it  is  occupied  on  but  a  day's 
treatment  of  only  a  circulating  by-product. 

On  the  other  hand,  the  dore-furnace  bottom  constitutes 
a  very  large  item,  due  to  the  facts  that  it  is  a  small  fur- 
nace which  means  disproportionate  brickwork,  and  that  it 
treats  something  more  than  the  entire  silver  and  gold  out- 
put in  six  day  charges.  The  net  result  is  that  this  accounts 
for  12  days  silver  and  10  days  gold. 

The  small  absorption  in  the  crucibles  is  neglected. 

Conductors. — We  next  have  the  copper  tied  up  in  con- 
ductors, etc,,  in  the  tank  house  and  the  silver  in  contact 
pieces  leading  the  current  in  and  out  of  the  cells  in  the  part- 
ing plant. 

These  copper  parts  in  the  tank  house  consist  of  busbars, 
cross  rods  (upon  which  the  cathodes  are  hung)  and  starting 
blanks  (the  plates  upon  which  the  starting  sheets  are 
deposited). 

It  is  customary  to  buy  the  main  busbars  outright  as  a 
charge  against  capital  expense,  so  these  do  not  enter  into 
this  discussion. 

The  other  small  parts  are  borrowed  from  the  copper  pass- 
ing through  the  plant  as  they  are  made  of  refined  copper 
and  can  be  melted  up  at  any  time  at  a  day's  notice,  were 
some  extraordinary  market  condition  to  call  for  a  final 
accounting  in  refined  copper.  Of  course  this  loan  has  to 
be  paid  for  in  " metal  interest"  as  it  correspondingly 


METALS  IN  PROCESS  37 

lengthens  the  time  required  for  refining.  This  method  is 
really  no  different  from  that  pursued  in  casting  lugs  upon 
the  anodes  in  order  conveniently  to  suspend  them  in  the 
tanks. 

The  items  in  question  amount  to  about  1.5  days  produc- 
tion of  copper. 

The  silver  contact  pieces  in  the  parting  plant  are  a  small 
item,  tying  up  about  0.1  of  a  day's  production. 

Finally  we  have  the  copper  molds  used  for  casting  the  wire 
bars,  ingots,  etc.,  and  these  tie  up  about  2.5  days  produc- 
tion. 

The  total  days  lost  from  " metals  withdrawn"  is  shown 
in  Table  9. 

TABLE  9. — METALS  WITHDRAWN 

Copper  Silver         Gold 

Tank  house  electrolyte 0 . 60 

Parting  plant  electrolyte .  0 . 40 

Anode  furnace  bottom 0.61  0 . 33           0 . 52 

Refining  furnace  bottoms 0 . 52 

Dore  furnace  bottoms 12.00         10.00 

Conductors,  etc 1 . 50  0.10 

Molds..  2.50 


Total  days 5.73         12.83         10.52 

We  can  now  summarize  the  results  of  this  review  and 
Table  10  shows  a  total  time  required  of  45  days  for  the  cop- 
per, 69  days  for  the  silver  and  68  days  for  the  gold. 

TABLE  10. — SUMMARY  OF  METALS  TIED  UP 

Copper  Silver  Gold 

Commercial 5.00          4.00  3.00 

Technical — Direct  process. 24 . 00  34 . 00  37 . QO 

Circulating 3.93  4.21  3.66 

Withdrawn 5.73  12.83  10.52 

Lost 

Balancing 6.67  14.01  13.54 


Total  days 45.33         69.05         67.72 

To  these  figures  must  be  added  any  delays  due  to  fires, 
strikes  and  other  "unforeseen"  causes,  and  it  must  further 
be  remembered  that  running  a  plant  at  part  capacity  runs 
up  the  "  metals  withdrawn "  and  some  other  items. 


CHAPTER  III 


TANK  RESISTANCE 

The  cost  of  power  is  always  a  considerable  and  usually 
a  major  item  in  any  metallurgical  process  based  upon  elec- 
trolysis. In  the  ideal  refining  cell  the  energy  consumption 
would  be  zero  and  the  quantity  of  metal  recovered  per 
kilowatt-hour  therefore  infinite,  as  the  energy  liberated  at 
the  anode  would  just  offset  that  required  at  the  cathode. 
In  practice  there  are  ahost  of  resistances  and  counter  electro- 
motive forces  to  be  overcome,  and  a  detailed  study  is 
necessary  in  order  to  understand  the  possibilities  of  im- 
provement in  any  given  case.  In  this  chapter  practice  in 
copper  refining  by  the  multiple  process  will  be  inquired 
into  as  an  example. 

TABLE  11. — CLASSIFICATION  OP  TANK  RESISTANCE 


Item 

Class 

Nature 

Busbar  joints 

Contact 

Ohmic 

Busbars 

Conductor 

Ohmic 

Anode  contact  

Contact 

Ohmic 

Anode  lug 

Conductor 

Ohmic 

Anode 

Conductor 

Ohmic 

Surface  phenomena  

Transfer 

Ohmic 

Solution  of  anode.  .  . 

Electrochemical 

e.m.f. 

Slimes 

Resistor 

Ohmic 

Electrolyte  

Conductor 

Ohmic 

Deposition  of  cathode 

Electrochemical 

e.m.f. 

Surface  phenomena 

Transfer 

Ohmic 

Cathode  

Conductor 

Ohmic 

Cathode  loops  

Conductor 

Ohmic 

Loop  contacts  .    . 

Contact 

Ohmic 

Rod  

Conductor 

Ohmic 

Rod  contact 

Contact 

Ohmic 

Busbars  
Busbar  joints  

Conductor 
Contact 

Ohmic 
Ohmic 

38 


TANK  RESISTANCE 


39 


The  pounds  of  copper  recovered  per  kw.-hr.  expended  at 
the  switchboard  depends  upon  the  current  efficiency,  the 
current  density  and  the  items,  both  real  and  apparent, 
which  make  up  tank  resistance.  The  last  is  the  subject 
of  our  present  inquiry. 


Contacts  > 


Contact 


Contact 


7 


Adsorption  Phenomena     |K      'i.       Adsorption  Phenomena 
Energy  of  Combination-^Slimes^^Energy  of  DtssociaUon 


C.E.M.F. 


C.E.M.F. 


Conductor 


.Conductor 


Contact 


FIG.  10. — Diagram  of  circuit. 

If  we  follow  the  course  of  the  current  from  the  positive 
pole  at  the  switchboard,  through  the  tank  house  and  back 
to  the  negative  pole,  we  shall  find  the  series  of  obstacles 
to  its  passage  shown  in  Table  11,  taking  but  a  single  tank  in 
circuit  for  an  example.  This  is  also  shown  diagrammatic- 
ally  in  Fig.  10. 

It  is  evident  that  in  actual  practice  we  have  the  items  in 
the  tank  proper  multiplied  by  the  number  of  tanks  in  a 
circuit,  or,  what  is  the  same  thing,  we  may  work  out  the 
resistance  per  tank,  apportioning  to  each  tank  its  share  of 
the  busbar  resistance. 


40 


COPPER  REFINING 


The  fact  that  for  a  circuit  of  a  single  tank  the  busbar  re- 
sistance would  be  inordinate,  has  led  to  putting  a  number 
of  tanks  in  series  and  then  to  a  study  of  the  relative  ar- 
rangement of  groups  of  tanks.  The  placing  of  additional 
tanks  in  series  merely  distributes  the  voltage  drop  in  the 
leads  between  the  switchboard  and  the  tank  house. 

It  has  been  generally  considered  good  practice  to  keep 
the  line  voltage  down  to  200  volts  or  less,  which  places  an 
upper  limit  of  about  600  tanks  in  series.  Actually  circuits 
seldom  carry  above  400  tanks,  and  this  is  sufficient  to  make 
the  incoming  leads  amount  to  but  3  or  4  per  cent  of  the  total 
voltage  drop. 


ffiffll 


^Insulated 


Insulated 


FIG.  11. — Various  tank  and  electrode  arrangements. 

Figure  11  shows  the  evolution  of  tank  connections.  A  is 
the  arrangement  used  in  the  first  small  installations.  B 
is  a  modification  employed  at  the  old  Anaconda  refinery, 
long  since  dismantled.  Here  the  parallel  conductors  re- 
sulted in  halving  the  contact  resistance  between  con- 
ductor-bars and  tanks  without  using  any  more  copper,  as 
each  bar  was  made  half-size  (C). 

Then  A  was  expanded  into  twin  tanks,  as  shown  in  D. 
This  results  in  halving  the  conductor-bars  required,  saving 


TANK  RESISTANCE  41 

copper  investment  as  well  as  voltage  drop.  Further,  as 
the  connections  between  twin  tanks  placed  individual 
anodes  and  cathodes  in  independent  pairs,  it  was  claimed 
that  a  short  circuit  between  electrodes  in  one  tank  was 
limited  in  its  damage  to  efficiency  by  the  resistance  in  series 
in  the  adjoining  tank. 

The  old  Anaconda  tank  C  partly  met  this  argument  in 
that  it  was  very  wide,  and  two  anodes  were  hung  side  by 
side  from  a  single  cross-ban  This  required  hanging  the 
anodes  by  hooks,  however,  and  this  in  turn  increased  the 
number  of  contacts. 

Then  came  the  Walker  system,  shown  at  E,  where  the 
idea  of  B  was  expanded  indefinitely,  it  being  found  feasible 
to  sacrifice  accessibility,  which  was  still  preserved  on  one 
side  of  each  tank  at  Z),  to  power  and  investment  saving. 
This  resulted  in  a  great  saving  in  conductor-bars,  and  has 
been  generally  adopted.  The  connecting  strips  shown 
at  B  were  dropped  and  a  small  triangular  bar  running  the 
length  of  the  tank  partition  substituted  without  appreciable 
loss  in  efficiency.  These  bars  are  very  small  in  cross-sec- 
tion, as  they  carry  but  half  the  current  flowing  through  a 
single  electrode  when  a  tank  is  in  normal  condition. 

Finally  we  have  at  F  a  further  extension  of  the  group 
idea,  which  h^Mfeen  proposed  by  several,  in  which  the  in- 
dividual tanks  are  merged  into  one  great  basin,  the  Walker 
formation  of  electrode  connections  being  virtually  main- 
tained, the  tank  partitions  and  triangular  bars  being  re- 
placed by  a  suitable  iron  beam  to  carry  the  load  of  the 
electrodes.  This  plan  has  received  but  a  limited  applica- 
tion, as  it  introduces  in  a  modified  form  some  of  the  dis- 
advantages of  the  series  system  due  to  higher  voltages 
without  compensating  gams.  It  would  greatly  decrease 
the  first  cost  of  a  tank  house,  however,  as  well  as  that  of 
tank  repairs. 

An  idea  of  the  magnitude  of  the  different  items  constitut- 
ing tank  resistance  may  be  obtained  from  Table  12,  which 
gives  the  results  of  an  analysis  I  made  a  good  many  years 
ago  of  a  tank  house  built  on  system  D.  It  must  be  under- 


42 


COPPER  REFINING 


stood  that  the  ohmic  value  given  for  counter  electromotive 
force  is  simply  the  apparent  equivalent  under  the  condi- 
tions of  operation,  and  further  that  in  order  to  reconcile 
this  resistance  exactly  with  volts  divided  by  amperes  the 
current  efficiency  corrected  for  chemical  corrosion  would 
have  to  be  allowed  for. 

Following  this  introduction,  and  keeping  in  mind  that 
it  is  possible  to  reduce  any  of  these  resistances  to  very  low 
values  by  changes  in  either  construction  or  methods  of 
operation,  we  shall  now  discuss  these  possibilities  item  by 
item.  We  have  in  general  (1)  carrying  the  current  to  and 
from  the  electrodes,  (2)  from  the  electrodes  to  the  elec- 
trolyte, and  (3)  across  the  electrolyte,  and  we  shall  re- 
group the  items  in  Table  11  in  this  fashion  ha  order  to 
avoid  repetition. 

TABLE  12. — ANALYSIS  OF  TANK  RESISTANCE 


Item 

Ohms 
per  tank 

Per  cent 
of  total 

A.   Electrolyte... 

0  0000444 

55  1 

B.   Metallic  conductors  

0  0000131 

16  2 

C.   Contacts  

0  0000113 

14  0 

D.  Counter  electromotive  force  
E.  Slimes,  etc.,  by  difference  .    ... 

0.0000040 
0  0000078 

5.0 

9  7 

Total 

0  0000806 

100  0 

A.        Electrolyte  .  .  . 

0  0000444 

55  1 

B.  a.  Leads  

0  0000024 

3  0 

b.  Conductor  bars  

0.0000085 

10  5 

c.  Anodes  .    . 

0  0000002 

0  25 

d.  Cathode  rods. 

0  0000010 

1  2 

e.  Cathodes 

0  0000008 

1  0 

f  .    Connection  strips  

0  0000002 

0  25 

C.  a.  Anode  contact 

0  0000026 

3  2 

b.  Cathode  loop  contact  

0.0000043 

5.3 

c.  Cathode  rod  contact  

0.0000044 

5  5 

D.        Counter  electromotive  force 

0  0000040 

5  0 

E.        Slimes,  etc.,  by  difference 

0  0000078 

9  7 

Total  

0.0000806 

100.0 

TANK  RESISTANCE 


43 


Conductors. — These  consist  of  the  leads  from  the  switch- 
board to  the  tanks,  the  connections  between  the  leads  and 
the  electrodes  and  the  electrodes  themselves.  As  the 
resistance  of  a  conductor  varies  directly  as  its  length  and 

Amperes  per  Square  Inch 
0        100      200      300       400      500      600       700  '  800       900      1000    1100     1200    1300 


FIG.  12. — Data  for  determining  most  economical  section  of  copper  conductors. 

inversely  as  its  cross-section,  while  its  first  cost  varies  as 
the  product  of  the  two,  we  have  from  both  points  of  view 
to  make  each  connection  as  short  as  possible,  while  the 
cross-section  involves  a  balance  between  first  cost,  the 


44  COPPER  REFINING 

cost  of  power  and  sometimes  strength.  Carrying  capacity 
does  not  enter  as  far  as  heating  goes,  as  the  other  factors 
place  this  far  on  the  safe  side. 

As  we  have  a  steady  full  load  twenty-four  hours  a  day, 
we  can  apply  directly  Thomson's  law  that  the  cheapest 
cross-section  will  be  that  for  which  the  interest  on  the  copper 
investment  just  equals  the  cost  of  the  power  lost  by  the 
voltage  drop.  The  corresponding  current  density  to  be 
chosen  for  the  conductors  will  vary  greatly  with  the  cost 
data  for  the  individual  case.  This  density  is  usually  in 
the  neighborhood  of  500  amp.  per  sq.  in.,  as  against  1000 
amp.  commonly  used  in  switchboard  work. 

Figure  12  shows  these  relations  graphically.  A  standard 
conductor  1000  ft.  long  and  1  sq.  in.  in  cross-section,  is 
taken,  and  two  sets  of  curves  superimposed,  the  first  being 
the  interest  charges  on  the  copper  in  dollars  per  year  for 
different  prices  per  pound  for  copper  and  various  rates  of 
interest,  and  the  second  being  the  power  loss  in  dollars 
per  year  for  different  currents  and  various  costs  per  kw.  hr. 

It  is  in  this  way  possible  to  select  the  basic  data  and  at 
once  equate  the  two  values. 

For  example,  suppose  we  take  copper  at  15  cents  a 
pound,  interest  at  10  per  cent,  and  a  kw.-hr.  at  %  cent, 
we  enter  the  diagram  at  the  bottom  of  the  15-cent  line,  and 
note  that  it  intersects  the  10  per  cent  line  at  $57  a  year 
as  our  interest  charge.  This  same  $57  a  year  line,  however, 
intersects  the  curve  for  %  cent  kw.-hr.  at  an  abscissa  corre- 
sponding to  390  amp.  and  as  the  conductor  has  1  sq.  in. 
area,  this  means  a  current  density  of  390  amp.  per  sq.  in. 
An  additional  line  on  the  diagram  tells  us  that  the  voltage 
drop  will  be  3.4  volts  per  1000  ft.  of  conductor.  If  we  have 
300  tanks  on  a  circuit  absorbing  0.32  volt  each,  and  the 
generator  is  500  ft.  away,  we  should  have  a  line  voltage  of 
100  volts  and  3.4  per  cent  loss  in  the  leads,  exclusive  of  any 
excess  drop  at  joints  in  the  bars. 

The  same  principle  applies  to  the  various  connections 
around  the  tanks,  always  remembering  to  figure  out  just 
what  current  each  individual  piece  is  carrying.  Here, 


TANK  RESISTANCE  45 

however,  the  element  of  strength  enters  in,  and  it  may  be 
necessary  to  make  a  cathode  rod,  for  example,  larger  than  is 
required  for  current  carrying  capacity,  in  order  to  obtain 
requisite  stiffness.  In  some  plants  copper-covered  iron  is 
employed  in  such  cases. 

In  the  case  of  the  electrodes  themselves,  the  body  of  the 
cathode  and  anode  are  of  ample  cross-section.  The 
cathode  loops,  however,  are  sometimes  overlooked.  Sup- 
pose we  have  a  cathode  3  feet  square,  operating  at  a  cur- 
rent density  (of  electrolysis)  of  20  amp.  per  sq.  ft.,  hung  by 
two  loops,  each  3  in.  wide,  cut  from  starting  sheets  0.02  in. 
thick.  We  shall  have  a  current  of  3  X  3  X  2  X  20,  or 
360  amp.  carried  by  a  conductor  4  X  3  X  0.02,  or  0.24  sq. 
in.  in  section,  giving  a  density  of  1500  amp.  per  sq.  in.,  or 
far  above  the  economical  range. 

The  anode  metal  will  be  of  low  conductivity,  but  the 
lug  through  which  the  current  enters  is  usually  of  ample 
size. 

Contacts. — It  has  been  shown  that  the  size  of  the  metallic 
conductors  is  specifically  determined  by  the  cost  of  power 
and  other  considerations.  In  the  case  of  contact  resistances 
we  have  no  desirable  value,  the  proper  course  being  to 
make  them  just  as  small  as  possible.  Experiments  show  a 
contact  resistance  to  be  truly  ohmic  in  character,  the  voltage 
across  a  given  contact  increasing  directly  as  the  current  is 
increased.  It  appears  to  be  due  primarily  to  adsorbed  air 
on  the  surfaces  in  contact,  and  secondarily  to  oxide  or 
other  foreign  matter.  Pressure  and  moisture  lower  the 
resistance. 

We  have  two  classes  of  contacts,  one  where  permanent 
joints  can  be  mechanically  made,  and  the  other  where 
temporary  gravity  joints  must  be  used.  It  is  customary 
on  sliding  joints,  as  in  the  case  of  switch  surfaces,  to  keep 
the  surface  current  density  down  to  about  50  amp.  per  sq. 
in.  In  the  case  of  busbars  this  figure  can  be  greatly  ex- 
ceeded. A  planed  joint  firmly  bolted  together  will  operate 
at  200  amp.  per  sq.  in.  without  showing  appreciable  volt- 
age drop — that  is,  1  millivolt. 


46 


COPPER  REFINING 


In  the  case  of  the  loose  gravity  joints  between  the  elec- 
trodes and  their  supports,  we  have  one  of  the  large  sources 
of  waste  power  in  the  system  which  can  be  attacked  along 
three  lines,  namely,  decreasing  the  number  of  contacts, 
increasing  the  acting  pressure  and  improving  the  condition 
of  the  surfaces. 

As  the  surest  way  of  cutting  down  the  resistance  at  a 
joint  is  to  eliminate  the  joint,  study  has  naturally  been 
directed  toward  securing  the  minimum  number  of  loose 
joints  in  series  compatible  with  efficient  handling  of  the 
electrodes.  The  various  forms  of  anode  and  cathode  sus- 
pension are  shown  in  Fig.  13. 


FIG.   13. — Various  forms  of  electrodes. 

The  anodes  suspended  by  hooks  have  one  single  and  two 
twin  contacts  equivalent  to  two  series  contacts;  those  with 
loops,  one  single  and  one  twin,  equivalent  to  one  and  a 
half ;  while  the  standard  anode,  with  cast  lugs,  has  but  the 
single  contact  at  the  conductor  or  triangular  bar. 

The  cathodes  have  gone  through  the  evolution  shown. 
The  first  example  gives  but  a  single  contact,  but  has  been 
abandoned  on  account  of  trouble  with  corrosion  of  the  rods 
in  spite  of  painting  at  the  solution  line.  The  others  give 


TANK  RESISTANCE  47 

one  and  a  half  contacts,  but  a  more  reliable  method  of 
suspension.  The  part  of  the  loop  beneath  the  surface  of 
the  liquor  does  not  matter,  as  this  joint  is  soon  covered  by 
the  deposited  copper. 

At  one  of  the  refineries  work  has  been  done  on  ehminating 
another  contact  by  doing  away  with  the  triangular  bars, 
allowing  the  cathode  rod  to  rest  directly  on  the  lug  of  an 
anode  in  the  adjoining  tank.  This  is  in  effect  returning 
to  the  connection  shown  at  D  in  Fig.  11,  but  eliminating 
the  connection  strip. 

The  minimum  number  of  loose  contacts  we  can  get  along 
with  is,  therefore,  one-half  entering  the  anode,  one-half  at 
the  cathode  loops,  and  one-half  leaving  the  cathode  rod,  or 
an  equivalent  of  one  and  a  half,  while  the  usual  practice  is 
one  anode  and  one  and  a  half  cathode,  electrodes  being 
alive  at  one  lug  only. 

The  pressure  exerted  upon  the  two  surfaces  has  a  great 
deal  to  do  with  the  resistance  shown  to  the  passage  of  the 
current.  For  example,  the  contact  resistance  between  the 
triangular  bar  and  the  anode  increases  day  by  day  as 
the  anode  dissolves  away,  and  one  of  the  loop  contacts  will 
show  almost  as  high  a  resistance  as  the  cathode  bar  contact, 
which  has  twice  the  current  but  twice  the  weight. 

Various  devices  have,  therefore,  been  tried  to  increase 
this  pressure,  notably  plugs  and  clamps.  At  one  plant 
holes  were  drilled  in  the  anode  lugs  and  these  were  reamed 
out  to  receive  a  tapered  copper  plug  connected  permanently 
by  a  short  cable  to  the  busbar.  This  is  of  no  advantage 
if  the  original  contact  is  kept  clean,  because  we  are  dealing 
with  such  a  low  order  of  resistances  that  the  short  cable 
will  have  too  high  a  resistance  to  be  of  much  service  as 
a  parallel  circuit.  Of  course,  it  does  prevent  excessive 
values  for  the  joint,  but  with  present-day  methods  of  tank 
inspection  these  are  not  allowed  to  occur. 

A  number  of  years  ago  the  writer  tried  out  thoroughly 
the  use  of  spring  clips  on  the  cathode  loops  and  rod  connec- 
tions. These  gave  very  encouraging  results,  cutting  out 
over  80  per  cent  of  the  resistance.  When  the  cost  of 


48  COPPER  REFINING 

renewing  the  clips  from  time  to  time,  the  labor  of  handling 
them  and  the  additional  hindrance  in  working  around  the 
tank  was  allowed  for,  some  of  this  margin  was  eaten  up. 
The  great  offset  to  such  a  plan,  however,  has  been  the 
improvement  in  keeping  contacts  clean,  taken  up  below. 

Much  has  been  done  in  the  lowering  of  contact  resistance 
by  shaping  one  of  the  members  as  a  wedge  which  will  bring 
a  heavy  unit  pressure  upon  the  other.  This  is  the  idea  in 
the  triangular  bar,  and  it  has  been  carried  further  in  the 
proposal  to  have  a  wedge  of  a  different  angle  from  a  cor- 
responding groove  which  would  result  in  crowding. 

Finally  we  have  the  very  important  matter  of  the  condi- 
tion of  the  two  surfaces.  Contact  resistances  develop 
considerable  heat,  and  this  means  that  any  drippings  of 
electrolyte  in  the  neighborhood  of  a  contact  will  soon  be 
converted  into  anhydrous  copper  sulphate  forming  a  coating 
which  effectually  prevents  the  existence  of  a  perfect  contact. 

The  first  principle  is,  therefore,  to  keep  the  contacts 
clean,  and  when  set  up  the  rods  and  under  surface  of  the 
anode  lugs  are  always  brightened  with  sandpaper  or  its 
equivalent. 

Various  means  have  been  tried  for  making  a  better 
union  physically.  Mercury  cups  are  not  practical  for 
many  reasons,  but  amalgamating  the  surfaces  is  possible. 
This  gives  excellent  results,  but  the  cost  of  the  amalgam 
used,  plus  that  of  the  labor  applying  it,  amounts  to  more 
than  the  saving. 

At  one  time  a  scheme  was  advanced  for  keeping  the  anode 
and  cathode  rod  contacts  wet  by  substituting  shallow  cop- 
per gutters  for  the  triangular  bars  and  allowing  water  to 
flow  therein.  This  also  gave  good  results  but  at  too  great 
an  expense  for  proper  maintenance.  Finally  it  was  found 
that  oiling  a  contact  after  shining  it  did  not  interfere  with 
the  contact  itself,  while  it  did  serve  to  keep  it  clean  for  a 
long  time,  and  cost  almost  nothing  to  apply. 

The  result  of  these  various  developments  has  been  to  cut 
in  half  the  values  for  contact  resistance  given  in  Table  12. 

It  must  not  be  thought  that  because  the  ohmic  values 


TANK  RESISTANCE  49 

are  very  small  the  financial  equivalents  are  likewise  so. 
Take  the  saving  of  one-half  of  0.0000113  ohms  per  tank 
just  spoken  of,  and  assume  power  at  %  cent  per  kw.-hr., 
10,000  amp.  on  a  circuit  and  1500  tanks  in  the  tank  house. 

_,  .  0.0000113^10,000X10,000  .  1Knn 
The  saving  will  be ^~  -X-  1000  -  X  1500 

or  848  kw.,  equivalent  to  848  X  24  X  0.005  or  $102  a  day. 

Transfer  Resistance. — We  come  now  to  the  transfer  of 
the  current  between  the  electrodes  and  the  electrolyte. 
This  is  a  field  which  is  very  difficult  properly  to  resolve  into 
the  several  component  factors.  In  true  refining  the  total 
value  is  not  very  great,  but  bad  conditions,  such  as  poor 
circulation  of  the  electrolyte,  foul  anodes,  etc.,  may  greatly 
increase  the  normal  value.  In  general,  we  have  to  deal 
with  counter  electro-motive  force,  the  ohmic  resistance  of 
an  adsorbed  gas  film  and  the  screening  effect  of  the  slimes. 

The  counter  electro-motive  force  is  the  opposing  voltage 
due  to  the  cell  acting  as  a  battery,  and  is  due  to  the  differ- 
ence in  composition  between  the  anode  and  the  cathode, 
and  to  the  differences  in  concentration  of  the  electrolyte 
around  the  two  electrodes. 

The  fiist  cause  is  small  in  its  effect  except  in  the  insoluble 
anode  tanks  used  for  controlling  the  copper  contents  of  the 
electrolyte  which  operate  at  about  six  times  the  voltage 
required  for  refining  cells. 

The  second  is  due  to  the  fact  that  the  circulation  of  the 
electrolyte,  which  must  be  gentle  in  order  to  avoid  stirring 
up  the  anode  slimes  and  thereby  contaminating  the  ca- 
thodes, is  not  sufficient  to  sweep  away  from  the  face  of  the 
anode  the  descending  layer  of  solution  rich  in  copper  sul- 
phate formed  by  the  electrolysis,  nor  the  corresponding 
lean  layer  which  rises  at  the  surface  of  the  cathode. 

This  forms  a  Cu — CuSO4 — Cu  concentration  cell  with  a 
small  electromotive  force  tending  to  equalize  the  differ- 
ences in  concentration  and  therefore  against  the  applied 
voltage. 

It  is  quite  easy  to  measure  these  two  effects  jointly  by 
taking  careful  current-voltage  readings  while  varying  the 


50  COPPER  REFINING 

current  over  a  range  not  great  enough  seriously  to  change 
the  conditions.  If  these  readings  are  plotted  and  a  straight 
line  drawn  through  the  points,  this  line  will  intersect  the 
voltage  base  line  at  the  value  of  the  counter  electro-motive 
force  on  the  circuit.  In  practice  this  amounts  to  from 
0.01  to  0.02  volt  per  tank. 

The  ohmic  resistance  of  what  is  doubtless  an  adsorbed 
gas  film  on  each  electrode  is  considerable  and  under  some 
conditions  may  become  enormous. 

If  we  explore  the  potential  gradient  between  anode  and 
cathode  we  shall  find  a  sudden  drop  as  we  leave  the  anode, 
a  gradual  slope  across  the  electrolyte  and  another  sudden 
drop  as  we  reach  the  cathode.  This  film  acts  as  a  true 
ohmic  resistance  and  has  a  temperature  coefficient. 

The  various  addition  agents  which  have  proved  of  such 
assistance  in  obtaining  smooth  deposits  act  markedly  on 
this  resistance.  A  moderate  dose  of  gelatin  in  the  elec- 
trolyte will  increase  the  overall  voltage  required  as  much  as 
40  per  cent. 

Then  we  have  the  screening  effect  of  a  curtain  of  poorly 
conducting  slimes  on  the  face  of  the  anode.  A  well- 
refined  high-grade  anode  makes  a  loose  gianular  slime  which 
offers  but  little  resistance  to  the  passage  of  the  current; 
less  favorable  conditions  result  in  a  thick  greasy  slime  that 
is  pierced  in  places  by  the  current  making  for  high  resis- 
tance and  irregular  solution  of  the  anode.  The  latter  con- 
dition often  results  in  what  are  sometimes  called  "  crazy 
tanks"  where  a  voltmeter  across  the  tank  will  give  no  con- 
stant reading  but  jumps  violently  back  and  forth  between 
a  normal  tank  voltage  and  one  about  three  times  as  great. 

The  value  of  the  resistance  due  to  adsorbed  gas  and 
slimes  can  only  be  obtained  by  difference  between  the  sum 
of  the  values  found  for  all  other  items  and  the  total  over- 
all voltage.  This  is  generally  about  10  per  cent  of  the  total 
resistance  in  circuit.  It  probably  lies  chiefly  in  the  gas 
film  and  further  study  may  discover  some  way  of  reducing 
this  factor. 

Electrolyte. — We  come  finally  to  the  resistance  of  the 
electrolyte  itself  and  this  brings  up  three  questions,  the  nee- 


TANK  RESISTANCE  51 

essary  distance  between  the  anode  and  the  cathode,  the  com- 
position of  the  electrolyte  and  its  temperature.  As  the 
electrolyte  comprises  over  half  the  total  resistance  in  cir- 
cuit, it  is  necessary  that  it  be  considered  in  detail.  | 

The  question  of  permissible  electrode  spacing  belongs 
under  the  heading  of  current  efficiency  which  we  are  not 
here  discussing.  It  also  depends  upon  the  advisable  age 
of  electrodes,  or  particularly  upon  how  many  crops  of 
cathodes  correspond  to  a  single  set  of  anodes  as  each  crop  will 
operate  on  a  wider  spacing  than  the  previous  one.  When 
three  or  more  crops  are  drawn  it  may  pay  to  respace  the  tank. 

Spacing  also  is  related  indirectly  to  current  density  as 
additional  density  greatly  increases  the  difficulty  of  working 
at  close  spacing.  The  foulness  of  the  anode  also  has  a 
bearing  as  voluminous  or  flocculent  slimes  demand  greater 
distance  in  the  interest  of  a  clean  cathode. 

The  average  thickness  of  liquid  column  has  been  grad- 
usally  reduced  from  2  in.  to  about  an  inch  and  a  quarter, 
due  largely  to  better  control  of  the  physical  character  of 
deposits  in  late  years.  During  the  same  period  current 
densities  have  increased  and  the  area  of  electrodes  has  been 
enlarged  so  that  the  full  value  of  the  improvement  in  de- 
posit has  been  apportioned  in  several  directions. 

The  composition  of  the  electrolyte  is  very  important. 
It  may  be  considered  to  be  made  up  of  sulphuric  acid, 
cupric  sulphate,  impurities  in  the  form  of  sulphates  or  more 
complex  compounds  such  as  arseniates,  etc.,  and  finally 
addition  agents. 

The  conductivity  depends  chiefly  upon  the  mobile  hydro- 
gen ions  from  the  dissociation  of  the  sulphuric  acid  and,  as 
would  be  expected,  increasing  the  free  acid  within  certain 
limits  markedly  lowers  the  resistance  of  the  electrolyte. 

There  are  limitations  imposed  by  two  difficulties.  As 
we  have  a  solution  of  mixed  sulphates  we  are  bound  by  the 
habits  of  isohydric  solutions,  the  amount  of  dissociated 
hydrogen  depending  upon  the  relative  concentration  of  the 
other  sulphates  as  well  as  upon  that  of  the  sulphuric  acid. 

The  other  limit  is  due  to  the  fact  that  too  great  a  hydro- 
gen concentration  affects  unfavorably  the  electrode  surface 


52  COPPER  REFINING 

phenomena  discussed  under  a  previous  heading.  The  net 
result  of  these  two  limitations  is  that  very  high  percentages 
of  free  acid  do  not  give  improved  over-all  results  and  it  is 
not  customary  to  carry  above  13  per  cent. 

These  same  arguments  require  carrying  as  small  an 
amount  of  copper  sulphate  in  solution  as  shall  give  a  satis- 
factory deposit  at  the  cathode.  This  is  borne  out  in 
practice  although  the  variation  in  conductivity  is  not  very 
great  with  changes  in  the  copper  concentration.  A  com- 
plete set  of  measurements  of  the  conductivity  of  different 
mixtures  of  copper  sulphate  and  sulphuric  acid  has  been 
made  by  Richardson  and  Taylor.1 

With  good  operating  conditions  the  copper  in  the  electro- 
lyte can  be  carried  down  below  2  per  cent  without  impairing 
the  cathode  deposit ;  were  it  possible  to  increase  the  circula- 
tion, even  lower  values  could  be  considered.  It  is  unwise  to 
proceed  too  far  in  this  direction,  however,  and  values 
between  2.5  and  3.0  per  cent  are  considered  good  practice. 
Even  under  difficult  operating  conditions  there  is  no  particu- 
lar advantage  in  carrying  over  3  per  cent  copper.  The 
early  electrolytes  were  carried  at  4  per  cent  copper  and  8 
per  cent  free  acid;  these  have  gradually  been  modified  to 
2.75  per  cent  copper  and  12  per  cent  free  acid. 

The  specific  resistance  of  such  an  electrolyte  at  120°F. 
will  be  about  0.7  ohm  per  cubic  inch.  The  various  im- 
purities in  the  electrolyte  will  increase  this  anywhere 
from  5  to  15  per  cent  so  that  a  working  value  will  be  about 
0.8  ohm.  A  reasonably  accurate  measurement  of  this 
resistance  may  be  obtained  with  an  ordinary  voltmeter  and 
two  copper  electrodes  if  the  column  of  electrolyte  measured 
be  long  enough  to  render  negligible  the  voltage  effects  at 
the  electrodes.  Whether  this  length  has  been  obtained 
may  be  tested  by  increasing  it  and  seeing  if  any  lower 
readings  per  unit  of  length  are  obtained.  Eighteen  inches 
between  electrodes  will  generally  be  found  sufficient. 

The  minute  quantities  of  organic  addition  agents  have 
probably  but  slight  effect  upon  the  conductivity  of  the 

1  Trans.  Am.  Electroch.  Soc.,  Vol.  xx,  p.  179. 


TANK  RESISTANCE  53 

electrolyte,  their  effect  upon  the  resistance  being  at  the 
electrodes.  On  the  other  hand,  inorganic  agents  such  as 
ammonium  sulphate,  used  to  be  added  in  large  quantities 
and  these  had,  of  course,  to  be  reckoned  with,  any  increase 
in  sulphates  tending  to  drive  back  the  dissociation  of 
hydrogen  ions. 

The  temperature  of  the  electrolyte  is  a  very  important 
matter.  In  the  first  place  the  electrolyte  itself  has  a 
positive  temperature  coefficient  of  about  0.5  per  cent  per 
°F.  There  is  not  only  this  enormous  premium  set  upon 
running  with  the  solution  hot,  but  in  addition  the  electrode 
conditions  are  greatly  benefited.  The  disadvantages  are 
the  cost  of  heating,  the  increased  humidity  of  the  atmos- 
phere in  the  tank  house  and  the  increased  growth  of  copper 
in  the  electrolyte  by  chemical  action. 

It  is  customary  to  heat  the  liquors  to  about  135°F.  in 
the  circulation  wells  and  this  temperature  drops  10  to  20 
deg.  in  passing  through  the  system,  resulting  in  different 
resistances  in  different  tanks. 

Some  of  the  older  plants  had  long  cascades  of  tanks,  the 
electrolyte  flowing  through  five  or  six  tanks  in  series; 
modern  plants  have  generally  but  two  tanks  in  series  so 
that  the  temperature  inequalities  are  not  so  severe  as 
formerly. 

We  have  now  discussed  item  by  item  the  various  com- 
ponents of  tank  resistance.  In  making  these  up  into  a 
sum  to  compare  with  the  readings  of  the  switchboard 
instruments  we  must  see  to  it  that  we  have  properly 
allowed  for  the  number  of  series-parallel  circuits  formed  by 
the  multitude  of  anode-cathode  pairs,  for  the  proportion  of 
tanks  which  are  "locked  out"  so  many  hours  a  day  for 
replacement  of  electrodes  and  cleaning  of  slimes,  for  the 
special  conditions  in  insoluble  anode  tanks  and  finally 
for  the  negative  factor  introduced  by  imperfect  cunent 
efficiency  which  provides  a  by-pass  or  parallel  circuit  for  a 
certain  part  of  the  current,  the  discussion  of  which  will  be 
taken  up  in  a  later  chapter. 


CHAPTER  IV 
CURRENT  DENSITY 

The  current  density,  or  amperes  per  square  foot  of 
active  cathode  surface,  is  the  factor  in  an  electrolytic 
process,  such  as  copper  refining,  upon  which  above  all 
others  the  design  and  operation  of  the  plant  depend,  as 
upon  it  hangs  a  string  of  minor  factors  which  must  be 
properly  correlated  in  order  to  obtain  the  best  return  upon 
the  investment.  These  factors  may  be  classified  under 
cost  per  pound  of  cathodes  recovered,  first  cost  of  plant,  and 
metallurgical  purity  of  product. 

Aside  from  commercial  considerations  there  is  a  limit 
to  the  density  which  can  be  employed  imposed  by  the 
temperature  of  the  electrolyte.  By  far  the  larger  part  of 
the  electrical  energy  called  for  is  converted  into  heat  in 
overcoming  the  ohmic  resistance  of  the  cells  and  the  tem- 
perature of  the  electrolyte  rises  until  the  heat  losses  offset 
the  C2R  gain. 

The  electrodes  are  spaced  so  closely  in  the  tanks  that 
the  equivalent  energy  to  be  dissipated  is  quite  large  and 
further  at  about  150°F.  a  liquid  will  begin  to  steam  suffi- 
ciently to  make  a  tank  room  too  foggy  for  comfort  or  for 
efficient  inspection.  As  a  starting  point  we  may  first 
examine  this  question  of  temperature  of  electrolyte. 

Temperature  of  Electrolyte. — In  ordinary  practice  with 
moderate  densities  the  electrical  energy  is  supplemented 
by  a  certain  amount  of  either  live  or  exhaust  steam  to 
obtain  the  desired  temperature  of  operation  generally  in 
the  neighborhood  of  130°F.  This  require^  a  balancing  of 

54 


CURRENT  DENSITY  55 

the  cost  of  heating  against  the  gains  resulting.  In  the  last 
chapter  we  have  considered  in  detail  the  question  of  tank 
resistance  and  took  as  an  example  a  certain  case  which 
gave  the  distribution  of  resistance  shown  in  Table  12. 

Increasing  the  temperature  of  the  electrolyte  will  lower 
its  resistance  by  an  amount  proportionate  to  the  increase 
and  to  its  temperature  coefficient  of  about  0.5  per  cent  per 
°F.  It  will  increase  the  resistance  of  the  various  metallic 
resistances  and  contacts  which  it  can  affect. 

As  the  anodes  and  cathodes  are  the  only  submerged 
metallic  resistances  and  as  they  are  of  negligible  resistance, 
we  may  turn  to  the  exposed  resistances  and  contacts.  The 
leads  and  conductor  bars  along  the  sides  of  the  tanks  are 
too  far  removed  from  the  source  of  heat  to  be  much  affected. 
The  contacts  are  kept  well  above  room  temperature  by 
their  own  C2R  loss. 

The  temperature  coefficient  of  copper  is  plus  0.24  per 
cent  per  °F. 

I  think  we  may  safely  assume  that  the  total  change  in 
these  minor  resistances  with  change  in  temperature  of  the 
electrolyte  is  very  small.  The  "transfer"  resistance 
at  the  surfaces  of  the  electrodes  is  markedly  lowered  by 
increase  of  temperature,  but  this  gain  is  large  only  in  the 
lower  part  of  the  temperature  range. 

We  may  therefore  consider  the  temperature  coefficient 
as  approximately  minus  0.5  per  cent  per  °F.  at  high  temper- 
atures and  a  larger  figure  at  low  temperatures  due  to  the 
added  change  in  " transfer"  resistance. 

When  running  at  about  sixteen  amperes  per  square  foot 
of  cathode  surface  without  the  use  of  any  steam  for  heating, 
the  electrolyte  in  an  average  system  will  run  about  25°F. 
above  the  temperature  of  the  atmosphere  in  the  tank  room, 
which  latter  is  generally  around  70°F.  and  very  humid  un- 
less an  adequate  system  of  forced  ventilation  is  employed. 

Some  simple  tests  where  a  circuit  was  allowed  to  run 
without  any  heating  steam  until  at  temperature  equilibrium 
and  then  heating  as  rapidly  as  possible  in  the  solution  wells 
showed  the  results  given  in  Table  13. 


56 


COPPER  REFINING 


TABLE  13. — EFFECTS  OF  TEMPERATURE  OF  ELECTROLYTE   ON   TANK 

RESISTANCE 


Test 

Ohms  per  tank 

Deg.  Fahr. 

Ohms 

Temp. 

No. 

Initial 

Final 

Diff. 

Initial 

Final 

Diff. 

per 

degree 

per  cent 

1 

0.000094 

0  .  000089 

0.000005 

102 

121 

19 

0.00000026 

0.29 

2 

0.000097 

0  .  000087 

0.000010 

103 

122 

19 

0.00000053 

0.61 

3 

0.000091 

0.000082 

0  .  000009 

108 

128 

20 

0  .  00000045 

0.55 

4 

0.000096 

0.000088 

0.000008 

105 

118 

13 

0  .  00000062 

0.70 

Av. 

0.000095 

0.000087 

0.000008 

105 

122 

17 

0.00000047 

0.54 

| 

Current  =  4400 
amperes 


Watts  per  tank  without  steam  =  1840 


20 


loooo 

9000 
8000 
7000 

6000 

,§  5000 

f 

f  4000 

£3000 


2000 


1500 


Degrees  Fahr.  above  Room  Temperature 

30  40  50  CO          70 


FIG.  14. — Watts  to  maintain  temperature  of  electrolyte. 


CURRENT  DENSITY  57 

This  temperature  coefficient  of  0.54  per  cent  per  degree 
Fahrenheit  is  so  large  that  it  at  once  becomes  apparent 
that  the  application  of  direct  heat  to  the  electrolyte  should 
be  considered.  Some  interesting  tests  were  carried  out  at 
another  plant  by  measuring  the  temperature  of  the  elec- 
trolyte, amperes  and  voltage  from  anode  ear  to  cathode 
loop,  for  different  current  densities  without  direct  heat. 

Some  of  the  results  obtained  are  shown  in  Fig.  14. 
The  watts  per  tank  required  to  maintain  a  certain  tempera- 
ture is  a  measure  of  the  rate  of  cooling  at  that  temperature 
and  the  curve  is  therefore  exponential.  Plotting  the 
data  on  logarithmic  section  paper  indicates  a  law  of 

Watts  per  tank  =  20.8  X  (°F.  liquor  - 
°F.  atmosphere)1-45. 

This  formula  is,  of  course,  only  good  for  the  conditions  at 
the  particular  tank  house  where  the  tests  were  conducted. 
At  a  room  temperature  of  70°F.  this  formula  would  give  our 
practical  limit  temperature  of  150°F.  in  the  electrolyte  with 
12,000  watts  per  tank. 

Under  the  conditions  of  this  set  of  tests  the  internal 
tank  resistance  would  be  about  0.00003  ohm  and  12,000 
watts  would  correspond  to  a  current  of  20,000  amp.  and  a 
current  density  of  64  amp.  per  square  foot.  This  is  far 
above  any  density  ever  likely  to  be  applied  in  practice  un- 
less marked  changes  in  the  general  arrangement  of  the 
process  develop. 

With  a  density  of  45  amp.  per  square  foot  the  electrolyte 
would  probably  run  about  120°F.  and  the  use  of  steam  for 
heating  would  hardly  be  justified.  At  lower  densities  we 
have  the  usual  case  of  heating  in  the  solution  wells  by  steam. 

We  shall  see  in  Chapter  XI  that  if  we  figure  out  a  heat 
balance  between  live  steam  for  heating  and  the  saving  at 
the  boilers  for  decreased  electrical  power  for  electrolysis,  the 
total  cost  increases  as  the  electrolyte  is  heated. 

The  earlier  plants  did  use  live  steam  for  heating  but  the 
more  recent  installations  have  used  exhaust  steam,  in 
some  cases  obtained  by  operating  some  of  the  main  generat- 


58 


COPPER  REFINING 


ing  units  under  partial  vacuum,   and  this   changes  the 
entire  situation. 

Another  result  of  heating  the  electrolyte  is  to  increase  the 
growth  of  copper  in  solution  due  to  the  oxidizing  effect  of 
the  liquor.  Copper  is  not  ordinarily  considered  soluble  in 
dilute  sulphuric  acid  but  when  the  solution  is  aerated  by 
passing  from  tank  to  tank  the  oxygen  dissolved  brings  about 
slow  solution. 


Temperature  of  Electrolyte,  °Fahr. 
50      60      70      80       90     100     110     120     130     140     150    160    170     180    190 


II 


z 


FIG.  15. — Chemical  corrosion  of  copper  by  electrolyte. 


Figure  15  shows  the  rate  of  action  of  a  solution  carrying 
3  per  cent,  copper  and  varying  percentages  of  free  acid  upon 
a  piece  of  cathode  copper  about  3X3  inches,  as  the  tem- 
perature is  raised  in  a  beaker.  This  shows  that  the  degree 
of  attack  rapidly  increases  with  rising  temperature, 
and  is  practically  independent  of  the  degree  of  acidity. 
This  action  may  or  may  not  be  desirable  according  to 
circumstances. 

Where  quite  pure  anodes  are  being  refined  the  excess 
copper  builds  up  in  the  electrolyte  and  is  usually  removed 
by  means  of  insoluble  anode  or  " liberating"  tanks.  As 
these  operate  at  from  2.0  to  2.3  volts  or  some  seven  times 


CURRENT  DENSITY  59 

as  much  as  ordinary  depositing  tanks  require,  this  power 
tends  to  offset  the  resistance  saving  due  to  hot  electrolyte. 

When,  on  the  other  hand,  the  anodes  carry  considerable 
percentages  of  nickel,  cobalt,  etc.,  a  proportion  of  the  cur- 
rent is  employed  dissolving  these  substances  electrochemi- 
cally  at  the  anode  while  depositing  an  equivalent  quantity 
of  copper  at  the  cathode,  so  that  the  nickel,  etc.,  will  grow 
in  the  electrolyte  which  is  at  the  same  time  depleted  in 
copper.  Then  the  addition  of  copper  by  chemical  corro- 
sion is  very  welcome;  in  fact  in  extreme  cases  this  effect 
has  to  be  artificially  increased  by  the  use  of  towers  of  shot 
copper  or  similar  means. 

In  general  it  is  found  necessary  to  use  from  2  per  cent, 
of  liberating  tanks  at  one  extreme  to  no  liberating  tanks  and 
shot  towers  at  the  other. 

The  temperature  of  the  electrolyte  also  has  a  bearing 
upon  the  shrinkage  in  volume  in  the  electrolyte  due  to 
evaporation.  Now  that  it  is  customary  to  use  a  purifying 
process  on  a  closed  cycle  evaporation  gives  almost  the 
only  means  of  making  room  for  water  used  in  rinsing  the 
anodes  and  cathodes. 

The  final  and  really  controlling  reason  for  the  use  of 
warm  electrolyte  is  the  greatly  improved  metallurgical 
conditions  resulting  therefrom.  Not  only  is  the  cathode 
smoother  and  denser,  but  the  conditions  at  the  dissolving 
surface  of  the  anode  are  greatly  benefited. 

The  increased  temperature  brings  about  a  local  circula- 
tion which  assists  in  removing  the  dense  solution  of  copper 
sulphate  from  the  face  of  the  anode  and  hi  preventing 
stratification  in  the  tank. 

When  current  is  passed  through  a  cold  cell  without  any 
circulation  of  the  electrolyte  conditions  soon  become  so  un- 
balanced that  the  electrodes  will  gas,  the  voltage  show 
violent  fluctuations  and  the  anode  slimes  be  stirred  up 
and  carried  hi  suspension,  thereby  fouling  the  cathode. 

When  the  electrolyte  is  systematically  circulated  these 
bad  effects  are  counteracted,  but  the  rate  of  circulation 
which  can  be  employed  is  limited  by  the  eventual  stirring 
up  of  the  slimes  mechanically. 


60  COPPER  REFINING 

There  is  in  turn  a  current  density  for  this  limiting  circu- 
lation which  begins  to  bring  about  a  return  of  the  undesir- 
able gassiDg  and  for  a  cold  liquor  and  an  impure  anode  this 
density  is  so  low  as  to  require  an  abnormally  large  plant 
investment. 

When  the  electrolyte  is  heated  this  density  limit  is 
greatly  raised  and  it  was  practice  even  in  the  very  early 
days  to  heat  the  electrolyte  somewhat — perhaps  to  110°F. — 
and  later  this  temperature  has  been  gradually  increased 
until  135°F.  or  even  higher  entering  the  tanks  is  not 
uncommon. 

In  series  system  plants  the  tanks  have  commonly  been 
lined  with  some  asphalt  composition  which  softens  with 
heat  and  this  has  made  it  impracticable  to  use  temperatures 
as  high  as  would  be  otherwise  desirable,  and  is  one  of  the 
reasons  that  series  processes  require  relatively  pure  anodes. 

In  general,  therefore,  (1)  the  temperature  of  the  electro- 
lyte should  be  carried  well  up  toward  the  practical  limit 
of  150°F.  in  order  to  be  able  to  employ  as  high  a  current 
density  as  may  be  desirable  from  the  point  of  view  of  cost 
of  operation;  (2)  unless  a  very  high  density  permitted  by 
very  unusual  conditions  is  employed,  heating  in  the  solu- 
tion well  will  be  required  to  maintain  this  temperature;  (3) 
with  proper  exhaust  steam  design  this  heating  will  be  finan- 
cially profitable. 

Power. — Assume  a  tank  resistance  of  0.0000766  ohm,  a 
counter  e.m.f.  of  0.010  volt  per  tank,  24  pairs  of  2  ft.  X  3  ft. 
anodes  and  cathodes,  making  288  sq.  ft.  of  surface  and  90 
per  cent  current  efficiency.  As  we  increase  the  current 
density  we  shall  have  the  conditions  shown  in  Table  14. 

This  table  shows  that  the  power  cost  per  pound  of  copper 
rises  nearly  in  proportion  to  the  current  density  employed. 

It  is  true  that  we  have  assumed  a  constant  tank  resistance 
due  to  maintaining  the  electrolyte  at  a  constant  temperature 
by  means  of  steam. 

As  the  current  density  is  raised  the  electrical  heat 
dissipated  in  the  tanks  will  increase,  as  previously  dis- 
cussed and  this  will  mean  that  less  heating  steam  will  be 


CURRENT  DENSITY  61 

TABLE   14. — RELATION   BETWEEN  CURRENT  DENSITY  AND   POWER  COST 


Amperes 
per 
sa   ft 

Amperes 

Volts 

Watts 

Lbs.              Watt- 
copper            hours 
per  24            per  24 

Watt- 
hours 
perlb. 

escf.  1  1/. 

hours              hours 

copper 

5 

1,440 

0.120 

173 

81 

4,200 

52 

10 

2,880 

0.231 

665 

162 

16,000 

99 

15             4,320 

0.341 

1,470 

243. 

35,300 

145 

20 

5,760 

0.451 

2,600 

324 

.    62,400 

193 

25 

7,200 

0.562 

4,050 

405 

97,200 

240 

30 

8,640 

0.672 

5,810 

-486 

139,400 

287 

35 

10,080 

0.782 

7,870 

567 

189,000 

334 

40 

11,520 

0.892 

10,280 

648 

246,200 

381 

required  and  this  saving  will  give  some  credit  to  be  applied 
against  the*power*cost. 

4perK.W.H. 

0         0.1        0.2        0.3        0.4        0.5       0.6        0.7        0.8        0.9        1.0 


FIG.  16. — Current  density  vs.  power  cost  in  practice. 

It  is  evident  that  the  cost  of  producing  a  kilowatt-hour 
will  have  a  great  deal  to  do  with  the  question;  in  fact  it  is 


62 


COPPER  REFINING 


possible  to  plot  the  power  cost  at  various  plants  against 
the  current  density  used  and  obtain  quite  a  smooth  curve. 
Figure  16  shows  this  general  relation,  but  individual  cases 
may  show  exceptions  for  special  reasons. 

First  Cost  of  Plant. — Assume  that  the  tank  house 
equipment  in  the  example  taken  amounts  to  an  investment 
of  $300  a  tank  and  that  the  power  plant  costs  $80  a  kilowatt. 
As  the  former  will  vary  with  the  output  of  cathodes  per 
tank  day  and  the  latter  with  the  kilowatt  demand,  both  of 
which  depend  upon  the  current  density,  we  can  take  the 
relation  determined  in  Table  14  and  ascertain  the  joint 
influence  of  these  two  factors,  which  has  been  done  in 
Table  15. 


TABLE  15. — RELATION  BETWEEN  CURRENT  DENSITY  AND  COST  OF  PLANT 


Power 

Lbs. 

Tank  in- 

Watt- 

Watt- 

plant 

Total  In- 

Current 
density 

copper 
per 
tank 

vestment 
per  Ib. 
copper 

hours 
per 
Ib 

days 
per 
Ib. 

invest- 
ment 
per  Ib. 

vestment 
per  Ib. 
copper 

day 

per  day 

copper 

copper 

copper 

per  day 

per  day 

5 

81 

$3.70 

52 

2.16 

$0.17 

$3.87 

10 

162 

1.85 

99 

4.12 

0.33 

2.18 

15 

243 

1.23 

145 

6.04 

0.48 

1.71 

20 

324 

0.92 

193 

8.06 

0.65 

1.57 

25 

405 

0.74 

240 

10.00 

0.80 

1.54 

30 

486 

0.62 

287 

11.90 

0.95 

1.57 

35 

567 

0.53 

334 

13.90 

1.11 

1.64  • 

40 

648 

0.46 

381 

15.80 

1.26 

1.72 

Labor  Cost. — The  main  labor  items  in  operating  a  tank 
house  are  for  inspection  work  to  maintain  current  efficiency 
and  for  inserting  and  drawing  the  electrodes  as  the  copper 
moves  through  the  process. 

The  higher  the  current  density  the  more  difficult  is  it  to 
maintain  the  efficieif cy  on  account  of  the  increased  number 
of  short  circuits  caused  by  rough  deposits  on  the  cathodes. 

After  the  circulation  of  the  electrolyte  has  been  raised  to 
the  highest  rate  which  will  not  stir  up  slimes  and  addition 


CURRENT  DENSITY 


63 


agents  have  been  added  to  control  the  character  of  the 
deposit  as  much  as  possible,  the  only  ways  to  improve  effi- 
ciency are  either  steadily  to  increase  the  inspection  labor 
or  to  decrease  the  life  of  the  cathodes. 

Actually  as  the  density  rises  the  efficiency  in  practice  is 
sacrificed  a  little,  the  labor  is  increased  and  the  age  of  the 
cathodes  decreased  until  a  balance  is  struck. 

The  cost  of  making  starting  sheets  in  the  multiple  pro- 
cess also  enters  if  the  weight  of  the  individual  cathodes  is 
changed.  We  shall  confine  ourselves  to  an  inquiry  into  the 
effect  of  age  of  electrodes  upon  the  labor  cost. 


10          20         30 


Days 
50          60          70 


90        100 


FIG.  17. — Current  density  vs.  age  of  cathodes  in  practice. 

Age  of  Electrode s. — The  age  of  the  anodes  and  the  pro- 
portion of  anode  scrap  made  have  but  little  connection  with 
the  current  density,  except  that  foul  anodes  which  give 
heavy  scrap  will  work  at  increasing  disadvantage  as  the 
density  is  increased. 


64 


COPPER  REFINING 


In  the  case  of  cathodes  we  may  observe  a  general  relation 
between  age  and  density  as  shown  by  the  practice  at  dif- 
ferent plants  and  given  in  Fig.  17.  The  increasing  knowl- 
edge of  the  value  of  addition  agents  in  recent  years  has 
tended  to  lessen  the  slope  of  this  curve,  however,  and  the 
day  of 'the  very  low  density  plant  has  apparently  passed, 
in  America,  at  least. 

Taking  this  curve  as  a  basis  and  assuming  that  a  start- 
ing sheet  costs  one  cent  to  make  and  that  it  costs  two  cents 
a  cathode  to  handle  sheets  in  and  cathodes  out  of  the  tanks, 
making  a  total  of  three  cents,  we  can  figure  as  in  Table  16 
the  cost  of  changing  age  with  density. 


TABLE  16. — RELATION  BETWEEN  CURRENT  DENSITY  AND  AGE  OF  CATHODES 


Current 
density 

Lb.  per 

tank 
day 

Lb.  per 
cathode 
day 

Cathode 
age  in 
days 

Lb.  per 
cathode 

Cathodes 
per  ton 

Total 
per  ton 

5 

81 

3.4 

100.0 

340 

5.9 

$0.18 

10 

162 

6.7 

37.0 

248 

8.1 

0.24 

15 

243 

10.1 

15.2 

154 

13.0 

0.39 

20 

324 

13.5 

9.6 

130 

15.3 

0.46 

25 

405 

16.9 

6.8 

115 

17.4 

0.5V 

30 

486 

20.2 

5.2 

105 

19.0 

0.52 

35 

567 

23.6 

4.0 

95 

21.1 

0.63 

40 

648 

27.0 

3.2 

86 

23.3 

0.70 

Metallurgical  Effects  of  Current  Density. — A  high  cur- 
rent density  will  always  give  trouble  at  the  anode  when 
treating  foul  material  and  usually  bring  about  abnormally 
high  voltage  conditions.  But  even  where  normally  pure 
anodes  are  refined  the  disturbance  of  the  slimes  is  greater  as 
the  density  is  raised,  due  partly  to  the  necessarily  some- 
what increased  mechanical  circulation  and  partly  to  local 
gassing  or  heat  disturbances  at  the  face  of  the  anode. 

This  relation  can  be  clearly  traced  in  the  purity  of  the 
cathode  and  while  it  is  ordinarily  not  sufficient  in  magni- 


CURRENT  DENSITY 


65 


tude  to  affect  the  commercial  purity  of  the  copper  as  such, 
it  can  be  clearly  felt  in  the  silver  and  gold  losses  carried 
away  in  the  cathodes. 

The  percentage  of  the  silver  and  gold  in  the  anodes  which 
is  found  in  the  cathodes  at  various  plants  is  plotted  against 
the  current  density  in  use  in  Fig.  18.  Where  the  anodes 
are  rich  in  values  this  factor  may  be  of  some  importance. 


Percentage  of  the  Silver  and  Gold  in  the  Anodes  Lost  in  the  Cathodes 
0      0.2     0.4     0.6     0.8     1.0     1.2     1.4      1.6      1.8     2.0    2.2     2.4     2. 


40 

t 

]TZ 

35 

^ 

.  — 

"o  30 

& 

.* 

^ 

^ 

^ 

S 

1 

X 

/ 

S  i*> 

/ 

/ 

I 

a  10 
5 

/ 

/ 

i 

0 

/ 

! 

FIG.  18. — Current  density  vs.  metal  losses  in  practice. 


Financing  Metals  in  Process. — In  Chapter  II  it  was 
shown  that  the  .copper,  silver  and  gold  are  tied  up  in  the 
tank  house  an  equivalent  of  25  days  in  ordinary  eastern 
practice,  which  means  a  current  density  of  about  20  am- 
peres per  square  foot.  It  is  evident  that  this  time  will 
vary  with  any  change  in  the  density,  and  for  simplicity's 
sake  we  shall  assume  that  this  variation  will  be  in  straight 
inverse  proportion,  although  it  is  quite  possible  to  change 
the  weight  of  the  anode  or  percentage  of  scrap  at  the  same 
time. 

Taking  copper  at  a  price  of  fifteen  cents  a  pound  and 
silver  and  gold  values  of  varying  amounts  and  interest  at 
6  per  cent  per  annum,  we  have  the  situation  shown  in 
Table  17. 


66 


COPPER  REFINING 


TABLE  17. — RELATION  BETWEEN  CURRENT  DENSITY  AND  METALS  TIED-UP 


Interest  charges  per  ton  at  6  per  cent  with 

Current 

Days  in 

silver  and  gold  per  ton  at 

density 

tank  house 

$25 

$50 

$75 

$100 

$200 

$300 

5 

100.0 

$5.42 

$5.83 

$6.25 

$6.67 

$8.33 

$10.00 

10 

50.0 

2.71 

2.92 

3.13 

3.34 

4.17 

5.00 

15 

33.3 

1.81 

1.95 

2.08 

2.22 

2.78 

3.33 

20 

25.0 

1.36 

1.46 

1.56 

1.67 

2.08 

2.50 

25 

20.0 

1.08 

1.17 

1.25 

1.33 

1.67 

2.00 

30 

16.7 

0.90 

0.97 

1.04 

1.11 

1.39 

1.67 

35 

14.3 

0.77 

0.83 

0.89 

0.95 

1.19 

1.43 

40 

12.5 

0.68 

0.73 

0.78 

0.83 

1.04 

1.25 

Summary. — The  data  in  the  previous  paragraphs  are 
general  in  character  and  do  not  strictly  represent  any  par- 
ticular plant,  but  they  serve  to  illustrate  the  general  rela- 
tions between  current  density  and  cost  of  operation  which 
have  to  be  considered  in  designing  an  electrolytic  plant. 

The  first  step  in  such  a  design  after  determining  the  de- 
sired capacity  of  the  proposed  plant  and  the  probable  cost 
of  power  per  kilowatt  hour  is  to  combine  the  data  given 
under  the  various  preceding  headings  and  determine  the 
approximate  current  density,  thereby  settling  the  size  of 
the  tank  house  and  power  plant. 

As  an  illustration  and  without  carrying  such  an  analysis 
into  the  details  necessary  in  actual  work  we  will  assume 
that  the  gain  from  released  heat  just  offsets  the  loss  from 
decreased  current  efficiency  as  the  density  is  raised;  that 
interest  and  depreciation  on  the  tank  house  and  power 
plant  shall  be  20  per  cent  per  annum;  that  the  silver  and 
gold  values  in  the  anodes  amount  to  $100  a  ton;  and  that 
we  are  interested  in  five  locations  where  the  probable  power 
cost  is  estimated  at  KoC.,  J4c->  Kc->  %c-  and  Ic.  per  kilo- 
watt-hour respectively,  and  other  conditions  are  equal. 

These  assumptions  are  combined  with  the  data  from  the 
several  tables  and  diagrams  in  Table  18. 


CURRENT  DENSITY 


67 


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68  COPPER  REFINING 

The  figures  worked  out  for  total  costs  are  not  the  actual 
costs  of  refining,  but  simply  totals  which  vary  as  the  actual 
costs  will  when  the  density  is  varied.  In  order  to  find  the 
density  which  will  give  the  minimum  cost  for  each  power 
cost  these  results  are  plotted  in  Fig.  19. 

In  Fig.  20  the  minima  so  determined  are  plotted  against 
the  current  density  and  we  have  established  the  general 
relation  already  shown  in  Fig.  16  between  current  density 
and  power  cost. 

It  will  be  observed  that  the  two  curves  for  theory  and 
practice  agree  well  except  at  the  lower  power  costs  where 
practice  runs  at  a  higher  current  density  than  would  be 
expected  from  our  discussion. 

This  is  due  largely  to  the  fact  that  although  a  certain 
density  may  show  minimum  cost  a  higher  density  may  show 
maximum  profits.  For  example  suppose  that  a  plant 
operates  with  power  costing  >^c.  per  kilowatt  hour  at  the 
density  for  lowest  cost  sho\vn  by  our  curve  in  Fig.  20  to  be 
18  amp.  per  square  foot,  and  that  the  profit  from  refining 
is  $3.00  per  ton.  If  now  the  density  be  pushed  up  to  25 
amp.  a  square  foot  the  cost  by  Fig.  19  will  be  increased 
about  $0.33  a  ton  and  the  profit  per  ton  be  cut  to  $2.67; 
but  the  increased  density  will  allow  the  plant  to  treat  38 
per  cent  more  copper  so  that  instead  of  $3.00  we  make  an 
equivalent  of  1.38  by  $2.67  or  about  $3.70,  with  less  than 
proportionate  increase  in  capital  expense. 

This  argument  is  fallacious  where  we  are  designing  a 
plant  for  a  certain  tonnage,  as  our  calculations  have  shown 
that  the  $0.33  excess  cost  would  pay  more  than  20  per 
cent  on  the  increased  investment  to  bring  the  density  back 
to  18  amp.  per  square  foot,  but  practically  a  plant  after 
being  put  into  operation  is  crowded  by  the  steady  annual 
growth  in  the  production  of  copper  and  we  may  say  is 
unable  to  keep  up  with  this  demand  for  capacity  by 
extensions. 

On  the  other  hand  where  power  is  expensive  this  process 
of  expansion  receives  a  much  earlier  check  than  where  it  is 


CURRENT  DENSITY 


69 


0        1        2 


0 
FIG. 


Dollars  per  Ton 
45678 


10      11       12      13      14 


19. — Current  density  vs.  variation  in  total  operating  costs. 


0         0 


1        0.2        0.3        0.4        0.5        0.6        0.7       0.8         0.9       1.0 


FIG.  20. — Current  density  vs.  power  cost — theory  vs.  practice. 


70  COPPER  REFINING 

cheap  and  this  tends  to  explain  at  least  part  of  the  differ- 
ence between  theory  and  practice. 

Then  too  at  very  low  power  costs  we  must  be  consider- 
ing water  power  installations  and  this  not  only  affects  the 
investment  factor  we  have  assumed  for  steam  plants,  but 
introduces  the  question  of  secondary  water  power  where 
excess  power  allowed  to  run  to  waste  over  a  dam  may  be 
utilized  at  exceedingly  low  rates.  Also  the  assumptions  of 
interest  rates,  price  of  copper  and  silver  and  gold  values 
have  to  be  corrected  for  each  practical  case. 


CHAPTER  V 
CURRENT  EFFICIENCY 

Current  efficiency  in  an  electrolytic  operation  is  the 
ratio  between  the  weight  of  product  obtained  per  ampere 
hour  and  that  called  for  by  Faraday's  law,  for  the  current 
used  in  the  desired  reaction.  It  has  therefore  much  to  do 
with  the  cost  of  operations,  but  in  this  connection  it  must 
be  considered  jointly  with  the  voltage  factor;  for  example, 
the  series  system  in  copper  refining  operates  at  a  lower 
current  efficiency  than  the  multiple  and  yet  yields  a 
greater  weight  of  cathode  per  kilowatt-hour  expended  in  the 
cell.  For  any  given  case,  however,  it  is  desirable  to  obtain 
as  high  a  current  efficiency  as  may  be  consistent  with  the 
cost  of  securing  it,  and  it  is  our  purpose  to  examine  this 
question  in  its  bearing  to  the  multiple  system  of  electrolytic 
copper  refining. 

The  percentage  which  we  obtain  by  dividing  the  weight 
in  grams  of  cathode  per  ampere  hour  by  1.186  is  less  than 
100  because  part  of  the  current  performs  no  electrolytic 
operation  on  account  of  leakage  or  short  circuits  between 
electrodes,  part  is  involved  hi  reactions  not  desired  and  a 
portion  of  the  cathode  itself  is  redissolved  chemically. 
It  is  even  possible  to  obtain  apparent  efficiencies  above  100 
per  cent  under  certain  conditions. 

In  order  to  secure  an  exact  correspondence  with  Fara- 
day's law  various  niceties  of  operation  must  be  observed 
and  a  great  deal  of  study  in  this  direction  has  been  applied 
by  those  interested  in  voltameter  measurements.  It  is 
quite  possible,  however,  with  care  and  dismissal  of  ques- 
tions of  labor  cost,  to  obtain  a  working  efficiency  on  a  large 
scale  of  99  per  cent.  Commercially  it  is  found,  however, 
that  "about  92  per  cent  is  as  high  a  figure  as  it  is  advisable 

71 


72  COPPER  REFINING 

to  insist  upon  and  at  some  plants  90  per  cent  or  even  88 
per  cent  is  considered  satisfactory.  Current  density  has 
a  direct  bearing  on  the  problem  in  that  a  high  density  greatly 
increases  the  difficulty  in  preventing  short  circuits  between 
the  electrodes. 

The  various  factors  involved  may  be  classified  as  follows : 

A.  Current  leakage : 

a.  To  ground. 

b.  Through  electrolyte. 

c.  Between  electrodes. 

B.  Reaction: 

a.  Deposition  of  impurities. 

b.  Gassing. 

c.  Valence. 

C.  Cathode  shrinkage: 

a.  Sulphatizing. 

b.  Ferric  salts. 

c.  Nodules,  etc. 

A.  CURRENT  LEAKAGE 

(a)  To  Ground. — The  insulation  resistance  of  the  cir- 
cuit from  the  ground  may  be  determined  by  noting  the 
reading  of  a  high-resistance  voltmeter  when  connected 
from  either  switch  terminal  on  the  live  circuit  to  the  ground 
by  a  simple  application  of  Ohm's  law.  The  current  flowing 
through  the  voltmeter  is  obtained  by  dividing  its  reading 
by  its  known  resistance;  the  total  resistance  in  the  leakage 
circuit  is  obtained  by  dividing  the  known  line  voltage  by 
this  current ;  the  insulation  resistance  is  found  by  subtract- 
ing the  known  voltmeter  resistance. 

It  is  evident  that  the  resistance  found  in  this  way  is  a 
measure  of  the  obstacles  in  the  path  of  the  current  from  the 
ground  back  through  improper  channels  to  the  other  leg  of 
the  circuit,  and  that  by  opening  the  main  circuit  at  pre- 
determined points  and  taking  repeated  measurements  a 
leakage  map  could  be  worked  out.  This,  however,  should 
not  be  necessary,  as  while  the  insulation  resistance  of  an 
electrolytic  circuit  will  always  be  low — a  usual  value  is  five 
ohms — the  actual  loss  of  effective  current  from  this  source 


CURRENT  EFFICIENCY  73 

is  'a  minor  matter  in  any  plant  where  the  foundation  piers 
are  properly  capped  with  glass  plates  and  tank  leaks 
kept  from  resulting  in  sulphate  crystals  climbing  around 
promiscuously. 

If  we  assume  that  a  circuit  has  a  total  resistance  of  five 
ohms  from  one  side  to  the  other  through  the  ground  and 
that  the  line  voltage  is  150,  this  leakage  will  be  but  30 
amp.  If  the  main  current  is  10,000  amp.-  and  the  leakage 
uniformly  distributed  so  that  it  robs  an  average  of  but  half 
the  tanks,  the  percentage  loss  will  be  but  0.15  per  cent. 

(b)  Through  Electrolyte. — There  are  three  ways  of 
determining  the  loss  of  effective  current  due  to  improper 
shunt  circuits  through  the  circulation  system:  by  direct 
measurement,  by  Ohm's  law  calculations  and  by  Faraday's 
law  calculations. 

The  direct  method  consists  of  placing  carefully  cali- 
brated ammeters  at  various  points  in  the  circuit  and  com- 
paring their  readings,  which  would  be  identical  except  for 
losses  under  (a)  and  (b).  Considerable  care  is  required 
to  avoid  errors  in  measurement  as  thermoelectric  effects 
may  creep  into  the  temporary  shunt  connections  and 
magnetic  errors,  due  to  strong  stray  field,  may  distort 
the  meter  readings  even  when  the  instruments  are  pro- 
tected by  iron  cases.  An  unprotected  portable  instrument 
may  even  be  permanently  thrown  out  of  adjustment  by 
exposure  of  its  permanent  magnets  to  the  action  of  the 
stray  field  which  exists  within  a  couple  of  feet  of  a  conductor 
carrying  10,000  amp. 

Another  method  is  to  open  the  circuit  in  the  center  and 
note  how  many  amperes  are  recorded  by  the  power-house 
ammeter  when  full  voltage  is  applied.  This  is  not  fair 
in  that  the  voltage  distribution  throughout  the  circuit  is 
not  normal. 

It  will  generally  be  found  that  the  tanks  at  the  far  end 
of  a  circuit  receive  3  or  4  per  cent  less  than  the  switchboard 
current. 

Ohm's  law  calculations  may  be  made  upon  the  liquid 
columns  of  electrolyte,  as  we  know  the  resistance  per  cubic 


74  COPPER  REFINING 

inch  of  the  liquor,  its  physical  dimensions  and  the  voltages 
operating.  Where  the  voltages  are  sufficiently  low  the 
lead  pipe  cannot  act  as  a  conductor,  as  when  the  current 
leaves  it  there  must  be  sufficient  voltage  to  decompose 
water.  When  such  voltages  occur  sections  of  hard  rubber 
or  of  rubber  hose  are  employed  to  break  the  continuity 
of  the  metal  path.  It  is  also  possible  to  take  fall  of  poten- 
tial readings  along  a  lead  pipe  and  figure  the  current  flowing 
therein  by  Ohm's  law  and  the  specific  resistance  of  hard 
lead. 

Finally,  we  know  that  wherever  the  current  enters 
the  piping  system  it  must  deposit  Faraday's  equivalent 
of  copper.  This  is  of  great  practical  assistance,  as  while 
it  gives  no  information  regarding  the  current  flowing  in  the 
liquid  itself,  it  does  bring  to  daily  attention  any  abnormal 
participation  of  the  conduit  system,  and  the  accumulation 
of  copper  trees  demands  early  attention  to  avoid  stopping 
up  the  pipes. 

The  coating  of  lead  sulphate  which  covers  all  tank  linings 
and  pipes  exposed  to  the  action  of  the  electrolyte  acts  as  an 
insulating  paint  of  much  value,  as  shown  by  the  low 
efficiencies  always  obtained  in  starting  up  a  new  installation 
of  bright  tanks.  A  final  insulating  joint  is  usually  effected 
where  the  electrolyte  leaves  a  tank  by  allowing  it  to  fall 
freely  into  the  launder  without  a  containing  pipe  for  some 
inches. 

Altogether  ample  means  exist  for  measuring  and  con- 
trolling this  source  of  loss. 

(c)  Between  Electrodes. — The  direct  touching  of  anode 
and  cathode  is  the  most  usual  cause  of  poor  efficiency. 
This  condition  can  be  brought  about  by  the  electrodes 
being  carelessly  spaced  in  the  tank,  by  the  curling  of  start- 
ing sheets,  by  falling  of  anode  scrap,  by  omission  of  elec- 
trode insulators,  by  electrodes  displaced  sideways  so  as  to 
touch  the  lead  lining  of  the  tank,  by  " treeing"  of  the 
cathode  deposit,  by  the  accumulation  of  an  excessive 
quantity  of  slimes  in  the  bottom  of  the  tank  or  by  the 
careless  leaving  of  tools  lying  on  top  of  the  electrodes. 


CURRENT  EFFICIENCY  75 

It  is  obvious  that  with  the  exception  of  "treeing"  and 
falling  anode  scrap  these  causes  may  be  removed  in  propor- 
tion to  the  amount  of  labor  and  inspection  applied.  Tree- 
ing involves  the  control  of  the  cathode  deposit  by  the 
choice  of  a  suitable  cathode  age  for  the  current  density 
employed,  adequate  circulation  of  the  electrolyte  and  the 
use  of  addition  agents. 

A  cathode  deposit  starts  as  a  fine  frosting,  and  with  a 
violent  circulation  this  builds  up  with  perfect  smoothness. 
It  is  not  possible  to  employ  even  a  rapid  circulation, 
however,  on  account  of  the  consequent  stirring  up  of 
anode  slimes,  and  any  mechanical  scouring  effect  is  there- 
fore lost.  On  the  other  hand,  the  circulation  must  be 
maintained  at  a  rate  sufficient  to  supply  copper  ions  at  the 
cathode  as  fast  as  demanded  by  the  current  density,  or 
other  ions  will  act  as  carriers  of  the  current  and  the  deposit 
will  become  rough  and  non-adherent.  Even  under  normal 
conditions  fine  needles  soon  spring  out  in  crystal  formation. 
The  moment  the  cathode  surface  becomes  roughened  the 
parts  nearest  the  anode  offer  the  path  of  least  resistance  to 
the  current  and  bad  soon  becomes  worse. 

The  function  of  addition  agents  is  first  to  round  off  these 
needles  into  blunt  nodules,  and  second  to  change  what  is 
called  " cocoa  matting"  structure  to  a  hard  compact 
deposit.  Ordinary  lubricating  oil  possesses  some  property 
which  effects  the  first  and  minute  quantities  of  glue  the 
second. 

The  falling  apart  of  scrap  anodes  results  from  incomplete 
refining  of  the  blister  copper  in  the  anode  furnace.  As 
the  condition  causes  an  abnormal  quantity  of  scrap  and 
leaks  from  pierced  tank  linings  as  well  as  low-current  effi- 
ciency the  remedy  lies  in  better  furnace  treatment. 

A  short  circuit  between  two  electrodes  is  not  as  serious 
as  at  first  appears  because  there  are  a  certain  resistance  due 
to  conductors  and  contacts  in  series  with  each  electrode 
and  a  number  of  parallel  circuits.  For  example,  if  we 
have  thirty  pairs  of  electrodes  and  the  series  resistance 
amounts  to  one-quarter  of  the  total  across  the  tank,  it  is 


76  COPPER  REFINING 

evident  that  a  reduction  of  75  per  cent  of  the  resistance 
in  one  of  the  thirty  parallel  circuits  will  but  make  it  equal 
to  four  normal  circuits  out  of  the  thirty  and  not  draw  any 
great  share  of  the  current.  Then  the  increased  current 
flow  heats  the  conductors  and  contacts  affected,  and  this 
increases  their  resistance.  Finally,  there  is  a  certain 
rebate  on  the  voltage  side  due  to  the  lowered  resistance  of 
the  tank  when  it  comes  to  pounds  of  cathode  per  kilowatt- 
hour. 

In  general  an  economic  balance  is  struck  when  about 
5  per  cent  of  the  current  is  rendered  non-effective  by  local 
short  circuits. 

B.  REACTION 

(a)  Deposition  of  Impurities. — In  the  ordinary  depositing 
tank  the  current   consumed  in  the  direct   deposition  of 
impurities  is   obviously  negligible,   as   appears  from  the 
great  purity  of  the  cathode;  in  fact,  it  is  an  open  question 
whether  any  of  the  traces  of  impurities  found  in  cathode 
copper  are  due  to  electrolytic  action.     Where  insoluble  or 
partly  soluble  anodes  are  used  and  higher  voltages  .obtain, 
this  item  becomes  measurable.     An  example  of  the  first 
case  is  where  arsenic  is  deposited  in  a  "liberator"  tank  in 
copper  refining,  and  of  the  second  where  iron  is  deposited 
in  a  copper-nickel  refinery. 

(b)  Gassing. — A  copper  cathode  rarely  shows  upon  anal- 
ysis above  99.95  per  cent  copper.     The  metallic  impurities 
may  total  0.02  per  cent,  still  leaving  some  unaccounted  for 
difference.     Part  of  this  is  included  electrolyte,  but  after 
all  allowances  are  made  it  seems  probable  that  hydrogen  is 
present  either  as  hydride  or  by  occlusion.     We  know  that 
some  addition  agents  harden  the  cathode,  and  that  this 
hardness  may  be  removed  by  annealing.     When  all  the 
copper  in  a  liquor  is  plated  out,  as  in  the  case  of  an  elec- 
trolytic assay,  the  gassing  does  not  begin  and  the  voltage 
rise  suddenly  upon  the  exhaustion  of  the  copper,  but  some 
gassing  starts  early  and  the  voltage  gradually  rises.     In  the 
same  way  local  conditions  at  the  cathode  due  to  the  mod- 


CURRENT  EFFICIENCY  77 

erate  circulation  employed  may  cause  the  separation  of  a 
certain  amount  of  hydrogen  at  the  cathode  in  a  normally 
operating  cell.  It  seems  probable,  therefore,  that  a  small 
proportion  of  the  current  may  be  diverted  into  depositing 
hydrogen  instead  of  copper,  and  even  a  minute  quantity  of 
hydrogen  will  account  for  a  measurable  current  on  account 
of  its  very  low  electro-chemical  equivalent.  For  example, 
0.03  per  cent  of  hydrogen  would  take  0.9  per  cent  of  the 
current.  We  do  not  know  how  much  of  a  source  of  loss 
this  condition  is  in  straight  copper  work,  but  in  nickel  de- 
position it  may  be  enormous,  free  hydrogen  appearing  in 
quantity  at  the  cathode. 

(c)  Valence. — The  efficiency  is  based  upon  the  reaction 

CuSO4  +  H2O  +  current  =  Cu  +  H2SO4  +  O 

where  copper  acts  as  a  divalent  metal.  It  is  evident  that 
should  the  copper  in  the  electrolyte  be  present  in  a  cuprous 
salt,  copper  might  be  precipitated  at  an  apparent  efficiency 
of  200  per  cent.  Cuprous  salts  are  undoubtedly  formed  to 
a  certain  extent  at  the  anode,  as  shown  by  the  cuprous 
chloride  and  metallic  dust  found  in  the  slimes  resulting 
from  such  a  reaction  as  Cu2SO4  =  CuSO4  -f  Cu,  although 
the  latter  is  partly  due  to  Cu2O  +  H2SO4  =  CuSO4  + 
H2O  +  Cu. 

Cuprous  sulphate  is  very  unstable  or  we  should  gladly 
use  it  as  the  basis  of  an  electrolyte;  and  it  seems  very  un- 
likely that  an  appreciable  amount  exists  at  the  cathode. 
The  cuprous  chloride  found  in  the  cathode  when  excess 
chlorides  are  allowed  to  accumulate  in  the  electrolyte  is 
doubtless  due  to  direct  reduction  by  the  cathode  as  CuCl2 
+  Cu  =  2CuCl. 

C.  CATHODE  SHRINKAGE 

(a)  Sulphatizing. — A  certain  amount  of  the  deposited 
copper  is  redissolved  by  the  electrolyte.  While  copper  is 
not  normally  soluble  in  dilute  sulphuric  acid,  the  oxygen 
dissolved  in  the  electrolyte  aids  in  a  slow  attack — Cu  -f- 
H2SO4  +  O  =  CuS04  +  H2O  probably  expresses  the  com- 


78  COPPER  REFINING 

plete  reaction.  As  would  be  expected  the  action  is  particu- 
larly marked  at  the  solution  line  and  various  expedients, 
such  as  painting  or  changing  the  solution  level,  have  to 
be  employed  in  order  to  prevent  the  cathode  loops  cutting 
through  at  the  solution  line. 

The  amount  of  this  chemical  action  is  indicated  by  the 
growth  in  the  copper  content  of  the  electrolyte  after  cor- 
recting for  anode  impurities  dissolving  electrolytically  and 
cuprous  oxide  in  the  anode  dissolving  chemically,  and  it  is 
found  to  increase  rapidly  with  increase  of  temperature  of 
the  electrolyte. 

A  fair  figure  is  about  2  per  cent  of  the  deposited  copper, 
and  if  we  assume  that  half  of  this  came  from  the  anodes 
we  have  an  apparent  loss  in  current  efficiency  of  1  per  cent. 

(b)  Ferric  Salts. — If  the  anodes  are  not  free  from  iron, 
converter  anodes,  for  instance,  and  the  resulting  ferrous  sul- 
phate is  allowed  to  accumulate  in  the  electrolyte,  there  is 
a  tendency,  increasing  with  concentration,  for  this  salt  to 
be  oxidized  at  the  anode— 2FeS04  +  H2SO4  +  O  =  Fe2- 
(SO4)3  +  H20.     This  is  particularly  the  case  in  insoluble 
anode  tanks  where  the  ferrous  sulphate  acts  as  a  true  de- 
polarizer.    This  ferric  salt  is  again  reduced,  either  elec- 
trically   at    the    cathode— Fe2(SO4)  3  +  2H  =  2FeSO4  + 
H2SO4 — or    by    the    anodes    and   cathodes    chemically — 
Fe2(SO4)3  +  Cu  =  2FeSO4  +  CuSO4.     In  either  case  there 
is  a  diversion  of  the  current  from  its  normal  work  and  a 
corresponding  loss  in  current  efficiency.     Under  bad  condi- 
tions, such  as  obtain  in  leaching  copper  ores,  this  loss  may 
become  very  serious,  but  in  straight  refining  work  it  should 
be  entirely  negligible. 

(c)  Nodules,  Etc. — There  is  a  certain  apparent  loss  in 
efficiency  due  to  mechanical  shrinkage  of  the  cathode  from 
nodules  falling  into  the  slimes,  chiefly  due  to  inspection 
work  on  the  tanks.     This  material  is  screened  out  of  the 
slimes  later,  but  is  too  contaminated  to  be  considered  legi- 
timate production.     Unless  the  deposit  is  rough  the  amount 
of  this  shrinkage  will  be  but  a  small  fraction  of  1  per  cent. 


CURRENT  EFFICIENCY  79 

SUMMARY 

The  nine  sources  of  efficiency  loss  are  always  present,  but 
any  or  all  of  them  can  be  kept  down  to  a  very  small  quan- 
tity. On  the  other  hand,  most  of  them  may  become  very 
serious  under  undesirable  conditions.  In  general  we  can 
say  that  entire  disregard  of  conditions  may  result  in  an  effi- 
ciency as  low  as  60  per  cent,  poor  work  85  per  cent,  good 
balanced  operating  92  per  cent,  and  efficiency  regardless  of 
expense  99  per  cent. 


CHAPTER  VI 


IMPURITIES 

Electrolytic  refining  gives  a  triple  separation  of  the  vari- 
ous impurities  in  a  blister  copper  anode,  distributed  in  the 
anode  slimes,  the  electrolyte  and  the  cathode.  The  chief 
object  of  the  process  is,  of  course,  to  make  a  pure  cathode, 
and  secondarily  to  keep  impurities  from  mounting  too 
high  in  the  electrolyte,  in  order  to  keep  the  cost  of  purifying 
the  electrolyte  within  reasonable  limits.  The  ideal  process 
would,  therefore,  send  all  of  the  impurities  into  the  anode 
slimes,  which  would  then  be  worked  up  into  various  by- 
products, and  the  electrolyte  would  stay  of  a  constant 
composition.  Practically  a  considerable  proportion  of  the 
impurities  dissolve  in  the  acid  sulphate  electrolyte,  and 
steps  have  to  be  taken  for  systematic  purification  of  the 
solution. 

Both  the  problem  and  the  means  of  dealing  with  it  have 
greatly  changed  in  the  last  twenty  years.  Two  decades 
ago,  when  electrolytic  refining  was  in  its  infancy,  there 
were  large  quantities  of  black  copper  to  be  treated,  and 
many  of  the  smelters  were  producing  pig  high  in  arsenic. 
An  idea  of  this  situation  may  be  gained  from  some  repre- 
sentative analyses  of  pig  copper  of  that  period,  given  in 
Table  19. 

TABLE  19. — SOME  PIG  COPPERS  ELECTROLYTICALLY  REFINED  IN  1901 


Brand 

Arsenic 

Anti- 

Iron 

Sulphur 

Lead 

Copper 

mony 

Argentine  

0.46 

0.40 

0.03 

0.15 

96.5 

Chicago  

0.30 

0.19 

0.22 

1.72 

96.0 

Germania 

2  92 

0  17 

0  01 

0  07 

95  0 

Philadelphia  

0.86 

0.08 

None 

Trace 

96.2 

80 


IMPURITIES 


81 


This  whole  situation  was  changed  when  the  strongly 
reducing  action  of  the  black  copper  furnace  and  the 
practically  neutral  atmosphere  of  the  copper-producing 
reverberatory  were  supplanted  by  the  strongly  oxidizing 
Bessemer  converter,  and  the  widespread  use  of  sulphur  as 
a  carrying  agent  for  copper  brought  nearly  the  whole  pro- 
duction through  a  process  which  efficiently  removed  most 
of  the  objectionable  impurities. 

On  the  other  hand,  some  imported  copper  pig  is  of  less 
than  standard  purity,  and  there  is  a  steadily  increasing 
production  of  secondary  copper  from  plants  treating  junk 
by  the  old  processes,  and  they  often  produce  very  foul 
pig.  Some  representative  analyses  of  these  various  classes 
are  given  in  Tables  20,  21  and  22. 


TABLE  20. — ANALYSES  o*  SOME  EXAMPLES  OF  SECONDARY  PIG 


Brand 

Arsenic 

Anti- 

Sulphur 

Iron 

Nickel 

Lead 

Copper 

A 

0.03 

0.94 

0.92 

0.35 

5.49 

85.0 

B 

0.03 

0.05 

1.02          0.36 

0.32 

2.68 

84.8 

C 

0.05 

0.19 

0.93 

0.47 

5.25 

82.4 

D 

0.05 

2.25 

Trace 

Tmce 

None 

12.79 

82.6 

E 

0.05 

0.41 

5.65 

1.58 

2.02 

F 

0.05 

2.63 

0.75 

5.53 

0.33 

2.36 

84.9 

TABLE  21. — ANALYSES  OP  SOME  FOREIGN  PIG  COPPER 


Country 

Arsenic 

Anti- 
mony 

Sul- 
phur 

Iron 

Nickel 

Lead 

Cop- 
per 

Chili  
Japan  

0.10 
0.18 

0.01 
0.12 

0.74 
0.39 

0.99 
0  03 

0.22 
0  06 

0.03 
0  83 

97.4 
97  6 

Peru  
Spain         .    .  . 

0.13 
0  11 

0.20 

0  69 

0.03 
1  03 

Trace 
0  02 

0  13 

97  7 

82  COPPER  REFINING 

TABLE  22. — ANALYSES  OF  NORMAL  BLISTER  COPPER 


Country 

Arsenic 

Anti- 
mony 

Sul- 
phur 

Iron 

Nickel 

Lead 

Cop- 
per 

Australia 

0  001 

0  Oil 

0  224 

0  029 

0  047 

0  003 

99  30 

Canada  

0  037 

0  045 

0  090 

0  050 

0.411 

0.002 

98.80 

Mexico 

0  017 

0  260 

0  040 

South  America 
United  States. 

0.007 
0.008 

0.006 

0.188 
0.045 

0.037 
0.034 

0.044 
0.037 

0.028 
0.007 

99.20 
99.30 

We  have,  therefore,  the  general  problems  of  keeping  the 
impurities  out  of  the  refined  copper  and  of  working  up  such 
of  the  impurities  fom  the  anode  slimes  and  the  electrolyte 
as  may  show  a  commercial  profit. 

The  early  refineries  had  much  trouble  with  even  the 
first  leg  of  this  proposition,  and  the  uncertainty  as  to 
the  chemical  purity  of  electrolytic  copper  produced  in  the 
early  days  had  much  to  do  with  the  premium  established  in 
favor  of  Lake  Copper.  Arsenic  was  the  chief  enemy,  and 
its  elimination  from  the  electrolyte  became  the  main  metal- 
lurgical problem  of  the  plant. 

This  situation  brought  about  the  development  of  a  by- 
product bluestone  plant,  fed  by  systematic  withdrawals 
from  the  electrolyte,  the  final  mother  liquors  being  pre- 
cipitated upon  iron  and  discarded. 

Then  the  flood  of  incoming  arsenic  abated,  and  in  some 
cases  nickel  became  the  major  impurity,  bringing  about  the 
development  of  the  by-product  nickel  sulphate  plant,  the 
mother  liquors  being  returned  to  the  electrolyte. 

In  general,  one  of  these-  two  methods  of  control  of  the 
composition  of  the  electrolyte  have  been  used,  and  they  will 
be  examined  more  or  less  in  detail  later  on. 

The  anode  slimes  were  at  first  cupelled  with  lead,  and 
only  the  silver  and  gold  recovered,  these  being  parted  by 
sulphuric  acid.  Later,  lead  practice  was  discarded,  except 
by  those  plants  operated  in  conjunction  with  a  lead  refin- 
ery, and  much  work  has  been  done  upon  the  recovery  of 


IMPURITIES  83 

selenium,  tellurium,  platinum,  palladium,  arsenic,  anti- 
mony, bismuth,  etc.  A  quite  complicated  pyro-metallurgy, 
followed  by  electrolytic  parting,  has  been  developed,  and 
much  research  devoted  to  the  possibilities  of  a  full  wet 
process. 

A  discussion  of  the  general  question  of  impurities,  there- 
fore, falls  under  six  main  headings,  namely:  (A)  sources, 
(B)  exits,  (C)  distribution,  (D)  chemical  requirements  of 
refined  copper,  (E)  recovery  of  soluble  impurities  from  the 
electrolyte,  and  (F)  recovery  of  insoluble  impurities  from 
the  anode  slimes. 

A.  SOURCES 

Apart  from  the  entering  pig  copper  there  are  as  many 
sources  of  impurities  as  there  are  supplies  entering  the 
process.  While,  of  course,  many  of  these  sources  are  quite 
negligible,  some  are  of  sufficient  magnitude  to  be  worthy 
of  consideration.  Among  these  are  fuel,  fluxes  and  acids. 

Fuel  enters  the  process  at  various  stages,  but  the  only 
place  where  it  is  of  account  in  this  discussion  is  in  the  melt- 
ing of  cathodes.  Here  the  various  products  of  combustion 
and  particularly  the  sulphur  have  to  be  reckoned  with,  as 
sulphur  is  one  of  the  principal  impurities  in  refined  copper. 
The  discussion  of  this  question  falls  more  normally  under  the 
head  of  chemical  impurities  in  refined  copper. 

Fluxes  include,  by  a  somewhat  liberal  definition  of  the 
word,  the  materials  of  which  the  furnaces  are  made  insofar 
as  they  enter  the  metallurgical  slags,  the  true  fluxes,  such 
as  limestone  and  pyrites  cinder  added  in  the  retreatment  of 
these  slags,  the  charcoal  or  other  carbonaceous  covering 
used  to  protect  molten  copper  from  undue  oxidation,  bone 
ash  or  similar  material  used  to  "butter"  the  molds,  and 
soda  nitre  and  soda  ash  used  as  fluxes  chiefly  in  the  treat- 
ment of  the  anode  slimes,  and  the  antimonial  lead  with 
which  the  tanks  are  lined. 

Under  acids  we  have  sulphuric,  nitric  and  hydrochloric, 
the  last  sometimes  as  sodium  chloride.  The  sulphuric  acid 
is  added  to  make  up  losses  of  free  acid  in  the  electrolyte  in 


84  COPPER  REFINING 

the  copper  electrolysis  and  the  nitric  similarly  in  the  part- 
ing plant.  The  chloride  is  added  to  the  copper  electrolyte 
chiefly  to  precipitate  antimony  as  oxychloride,  although 
it  has  been  claimed  by  many  to  have  positive  value  as  an 
addition  agent. 

Fortunately  most  of  these  process  supplies  carry  but  little 
in  the  way  of  metallurgical  impurities.  The  notable  ex- 
ceptions are  blast-furnace  fluxes,  sodium  salts  and  sulphuric 
acid. 

As  refinery  slags  are  made  by  a  union  of  metallic  bases 
with  siliceous  furnace  material,  it  is  necessary  to  find  lime 
and  iron  to  replace  the  copper  in  these  slags.  The  natural 
source  of  iron  in  the  vicinity  of  most  refineries  is  pyrites 
cinder  resulting  from  the  manufacture  of  sulphuric  acid. 
As  arsenic  is  commonly  associated  to  a  greater  or  less  de- 
gree with  iron  pyrites,  more  or  less  of  this  element  may  be 
introduced  in  this  manner. 

In  the  same  way  sulphuric  acid  made  by  burning  pyrites 
is  likely  to  contain  considerable  quantities  of  arsenic  as 
an  impurity,  and  unless  purified  acid  is  used  it  may  be  a 
heavy  contaminating  agency,  since  where  sulphate  salts 
are  made  as  a  by-product  the  acid  purchases  are  large. 

In  like  manner  any  nitrate  of  soda  used  in  the  silver 
building  boiling  tanks  as  an  oxidizing  agent  or  sodium  chlo- 
ride added  to  the  electrolyte  directly  bring  about  a  con- 
centration of  sodium  sulphate  which  may  be  objectionable. 

B.  EXITS 

A  complete  analysis  of  what  may  be  called  " exits"  for 
impurities  has  been  given  in  Chapter  I.  Based  on  this, 
eliminating  such  items  as  do  not  bear  directly  on  our  im- 
mediate problem,  we  may  classify  the  exits  as  (a)  outgoing 
commercial  products,  (6)  slags  and  (c)  stack  gases. 

The  ideal  process  would  eliminate  the  two  latter  prod- 
ucts entirely  and  send  all  of  the  impurities  out  as  commer- 
cial products.  Practically  not  only  do  several  elements 
escape  in  their  entirety,  but  excepting  copper,  silver  and 


IMPURITIES 


85 


gold,  the  remaining  are  recovered  at  far  from  100  per  cent 
efficiency. 

The  tendency  to-day  is  to  check  stack  losses  so  that  in 
time  there  will  be  but  two  outlets,  the  one  to  commerce 
and  the  other  to  the  slag  dump. 

When  markets  are  unsatisfactory  some  of  the  products 
can  be  stored,  as  has  already  been  the  case  with  selenium, 
tellurium  and  bismuth. 

C.  DISTRIBUTION 

The  distribution  of  impurities  depends  partly  upon  their 
chemical  characteristics  and  partly  upon  the  metallurgical 
practice  of  the  individual  plant. 

In  the  first  place,  the  cathode  will  contain  measurable 
quantities  of  all  impurities  found  in  the  anodes,  although 
there  is  room  for  some  discussion  as  to  what  proportions  of 
these  arrive  by  electrolytic  deposition,  by  inclusion  of  elec- 
trolyte and  by  mechanical  contamination  by  anode  slimes.1 
The  percentage  of  anpde  impurities  found  in  the  refined 
copper  may  be  seen  by  direct  comparison  of  average  analy- 
ses over  a  long  period  of  operation,  as  in  Table  23. 

TABLE  23. — METALLURGICAL  EFFICIENCY  OF  REFINING 


Element                       Anode 

Wirebar 

Per  cent 
of  original 
impurity 

Efficiency 
of 
refining 

Copper  99  .  030 

99  939 

Silver        0  1687 

0  00131 

0  78 

99  22 

Gold                                     .        0  0051 

0  000013 

0  25 

99  75 

Sulphur  0.0075 

0.0029 

38  60 

61  40 

Nickel  ...      0.3200 
Lead                             ...      0  0567 

0.0037 
0  0020 

1.15 
3  52 

98.85 
96  48 

Arsenic                                         0  0523 

0  0015 

2  87 

97  13 

Antimony  0.0409 

0.0034 

8  32 

91  68 

Bismuth  0.0051 

Trace 

Tellurium  0.0282 
Selenium              .                 .        0  0682 

0.00015 
0  00040 

0.53 
0  59 

99.47 
99  41 

Iron  i     0.0181 

0.0030 

21  55 

78  45 

Addicks:  Trans.  Am.  Electrochem.  Soc.,  vol.  xxvi,  p.  51. 


86  COPPER  REFINING 

The  four  elements  showing  a  cathode  recovery  of  over 
99  per  cent  are  silver,  gold,  selenium  and  tellurium,  none 
of  which  dissolve  in  the  electrolyte.  Therefore,  the  me- 
chanical contamination  by  anode  slimes  is  less  than  1  per 
cent.  Then  we  have  nickel  at  1.15  per  cent;  this  element, 
while  present  in  oxidized  form  in  the  slimes,  goes  chiefly 
into  solution  as  sulphate.  The  same  is  true  of  arsenic 
which,  however,  forms  a  light  slime  which  readily  attaches 
itself  to  the  cathode.  Lead  comes  not  only  from  the  anodes 
but  from  the  tank  linings,  so  that  the  efficiency  is  not  quite 
true  in  this  case.  The  same  is  true  in  lesser  degree  of  anti- 
mony, as  hard  lead  is  universally  used  to-day  for  tank 
linings.  Antimony  is  further  precipitated  from  the  elec- 
trolyte as  oxychloride,  as  previously  described,  entering 
the  float  slime.  A  characteristic  analysis  of  this  sediment 
is  given  in  Table  24. 

TABLE  24. — ANALYSIS  OF  FLOAT  SLIME 

Element  Per  cent 

Copper 3.0 

Arsenic 13.0 

Antimony 30 . 0 

Bismuth 8.0 

Silver 4.0 

Iron..  0.3 


When  we  come  to  consider  the  efficiency  of  refining  with 
regard  to  iron  and  sulphur  we  must  remember  that  both 
these  elements  are  introduced  in  the  melting  of  the 
cathodes,  the  former  in  the  rabbles  and  tools  used  and  the 
latter  in  the  fuel  and  ladle  charcoal,  and  we  also  have  sul- 
phate sulphur  from  the  electrolyte,  so  that  the  figures  are 
misleading. 

We  may  say  in  a  general  way,  therefore,  that  the  effi- 
ciency of  refining  is  very  high  and  that  the  cathode  copper 
.offers  a  very  small  outlet  for  anode  impurities ;  further,  that 
the  great  bulk  of  those  impurities  which  are  soluble  in  dilute 
sulphuric  acid  will  concentrate  in  the  electrolyte. 


IMPURITIES 


87 


The  balance  of  the  impurities  must  go  into  the  slimes  and 
we  can  attempt  a  measure  of  this  by  comparing  the  assays 
of  the  anodes  and  of  the  raw  slimes  as  they  come  from  the 
tanks  except  for  boiling  free  of  soluble  salts.  This  is  done 
in  Table  25. 

TABLE  25. — CONCENTRATION  OF  ANODE  IMPURITIES  IN  SLIMES 


Element 

Anode, 
per  cent 

Slimes, 
per  cent 

Per  cent 
anodes 
X84.4 

Per  cent 
recovered 

Cooper 

98  14 

14  3 

Silver  
Gold  
Nickel  
Arsenic  
Antimony                        .    . 

0.417 
0.00711 
0.314 
0.236 
0  0906 

35.0 
0.643 
5.25 
2.68 
5.35 

35.2 
0.600 
26.5 
19.9 
7.6  - 

99.5 
(107) 
19.8 
13.5 
70.0 

Bismuth  
Sulphur  
Iron                                  .    .  . 

0.0088 
0.0037 
0  0123 

0.46 
1.69 
0.17 

0.74 
0.31 
1.04 

61.9 

(541) 
16.4 

Lead  
Selenium  
Tellurium  
Zinc  
Insoluble  

0.0456 
0.0479 
0.0318 
0.0100 
0.1213 

2.44 
5.70 
2.69 
Trace 
6.60 

3.85 
4.04 
2.68 
0.84 
10.2 

63.4 

(141) 
(100) 
None 
64.5 

The  last  two  columns  of  Table  25  assume  that  99.5  per 
cent  of  the  silver  was  in  the  slimes  the  remaining  0.5  per 
cent  being  in  the  cathodes.  On  this  basis  the  recovery  in 
the  slimes  of  the  remaining  elements  has  been  calculated. 
Of  course,  the  gold,  selenium  and  tellurium  should  also 
show  a  recovery  of  about  99.5  per  cent,  and  the  discrepan- 
cies simply  indicate  that  the  correspondence  between  iden- 
tity of  slimes  and  anodes  is  not  quite  exact. 

The  sulphur  shows  a  large  excess  due  to  the  fact  that  the 
sulphur  in  the  sulphuric  acid  of  the  electrolyte  has  com- 
bined with  some  of  the  impurities  as  sulphates  which  have 
not  proved  readily  soluble. 

Nickel,  iron,  zinc  and  arsenic,  as  would  be  expected, 
show  small  recoveries;  the  first  three  form  readily  soluble 


88  COPPER  REFINING 

sulphates  and  arsenious  acid  is  quite  soluble  in  the  elec- 
trolyte. Nevertheless,  with  the  exception  of  zinc  the 
slimes  retain  some  of  even  these  elements. 

Lead  which  has  a  but  slightly  soluble  sulphate,  anti- 
mony, which  is  precipitated  as  oxychloride,  bismuth  and 
siliceous  matter  show  high  but  not  perfect  recoveries. 

Much  of  this  group  goes  into  the  float  slime  which  is 
largely  separated  out  before  the  heavy  slimes  are  sent  to 
the  silver  building  and  this,  if  corrected  for,  would  make  a 
nearly  complete  slime  recovery. 

An  analysis  of  the  electrolyte  corresponding  to  the 
example  given  above  would  be  about  as  stated  in 
Table  26. 

TABLE  26. — REPRESENTATIVE  ANALYSIS  OF  ELECTROLYTE 

Specific  gravity 1 . 226 

Per  cent  free  acid 12.03 

Per  cent  copper 2 . 94 

Per  cent  nickel 1 . 48 

Per  cent  chlorine 0.0081 

Per  cent  arsenic .  .  . 0 . 916 

Per  cent  antimony 0 . 0350 

Per  cent  iron 0.060 

Per  cent  bismath 0. 0026 

Per  cent  zinc 0 . 0166 

Per  cent  alumina 0 . 0595 

Per  cent  calcium  sulphate 0. 1348 

Per  cent  magnesium  sulphate 0 . 0370 

Per  cent  sodium  sulphate '. 0. 5048 


A  comparison  of  the  relative  values  of  the  impurities 
in  the  electrolyte  with  those  in  the  anodes  confirms  in  a 
general  way  the  distribution  already  shown.  It  is  evident 
that  in  this  particular  case  nickel  is  the  controlling  impurity 
and  that  any  system  of  purification  of  the  electrolyte 
which  will  hold  this  element  at  the  desired  concentration 
will  automatically  take  care  of  arsenic  and  other  impurities. 

That  the  degree  of  concentration  of  soluble  impurities  in 
the  electrolyte  has  a  direct  bearing  upon  the  purity  of  the 


IMPURITIES 


89 


cathode  product  may  readily  be  proved  in  the  oases  of 
arsenic  and  nickel.  Tables  27  and  28  give  comparisons 
of  wirebar  content  of  arsenic  and  nickel  with  corresponding 


0.0045 
00040 

5* 

X 

n  oo^*\ 

X 

X 

? 

X 

0  0025 

X 

A 

X 

X 

c 

0  0015 

-M 

*£* 

A.    Lx 
y\ 

-*\B 

^ 

0.0010 

2t_ 





.__ 

— 





— 

0        0.4      0.8       1.2       1.6       2.0       2.4       2.8      3.2      3.6       4.0      4.4 

Percent  Arsenic  in  Electrolyte 
FIG.  21. — Arsenic  in  electrolyte  vs.  arsenic  in  wirebars. 


0.0070 
0.0060 

00 

I   0.0050 

e   0.0040 

.5 

1   0.0030 

z 

|   0-OC20 
*   0.0010 

< 

i 

\ 

,/ 

\  ^ 

x 

x^ 

-k; 

\ 

^x 

5 

\ 

V 

X 

^ 

X 

K 

xl 

X 

\ 

<X 

X 

)        0.2      0.4      0.6       0.8       1.0       1.2      1.4       1.6      1.8       2.0      2. 

Percent  Nickel  in  Electrolyte 
FIG.  22. — Nickel  in  electrolyte  vs.  nickel  in  wirebars. 


analyses  of  electrolyte  at  a  number  of  plants.  These  data 
are  plotted  in  Figs.  21  and  22.  In  the  case  of  arsenic  the 
line  cuts  the  zero  ordinate  at  0.0013  per  cent  indicating  this 
quantity  due  to  contamination  by  slimes. 


90 


COPPER  REFINING 

TABLE  27 


Plant 

Representing 

Per  cent  As  in 
electrolyte 

Per  cent  As  in 
wirebars 

A 

1909 

0.728 

0.0020 

A 

1910 

0.789 

0.0016 

A 

1911 

0.506 

0.0018 

A 

1912 

0.219 

0.0019 

A 

1913 

0.194 

0.0015 

B 

21  Months 

1.04 

0.0016 

C 

1  Lot 

0.25 

0.0016 

C 

5  Lots 

1.64. 

0.0025 

C 

4  Lots 

2.37 

0.0033 

C 

3  Lots 

3.55 

0.0041 

D 

25  Months 

1.34 

0.0018 

TABLE  28 


Plant 

Representing 

Per  cent  Ni  in 
electrolyte 

Per  cent  Ni  in 
wirebars 

A 

1909 

.95 

0.0041 

A 

1910 

.60 

0.0055 

A 

1911 

.73 

0.0066 

A 

1912 

.32 

0.0038 

A 

1913 

.78 

0.0037 

B 

10  Months 

0.973 

0.0045 

C 

Rough  figures 

1.00 

0.0016 

D 

1  Lot 

0.594 

0.0014 

D.  THE  REQUIREMENTS  OF  REFINED  COPPER 

This  subject  is  of  sufficient  importance  to  be  reserved  for 
treatment  at  length  in  Chapter  IX. 

E.  PURIFICATION  OF  THE  ELECTROLYTE 

Any  system  of  purification  must  regularly  withdraw 
sufficient  electrolyte  to  control  the  amount  of  the  chief 
impurity.  For  example,  in  the  analysis  given  in  Table  26 


IMPURITIES 


91 


nickel  is  the  element  which  would  first  grow  to  undesirable 
concentration,  although  arsenic  is  a  close  second  and  the 
maximum  allowable  values  of  various  elements  depend 
upon  conditions  under  which  an  individual  plant  is  operat- 
ing. Both  nickel  and  arsenic  have  been  allowed  to  reach 
3  per  cent  in  the  electrolyte  without  disaster  in  certain 
local  and  temporary  cases.  The  first  thing  is  therefore  to 
ascertain  what  quantity  of  electrolyte  must  be  daily 
withdrawn. 

I.  Purification  by  Cementation  upon  Iron. — The  early 
methods  of  purification  consisted  simply  in  cementing  the 
copper  upon  iron  and  throwing  the  copper-free  liquor 
away.  The  scrap  iron  would  be  piled  in  a  lead-lined  tank, 
the  liquor  run  in  and  brought  to  a  boil  by  heating  with 
steam  and  at  the  end  of  an  hour  a  bright  iron  rod  would 
not  tarnish  when  introduced,  indicating  complete  pre- 
cipitation of  the  copper.  The  liquor  was  then  run  to  the 
sewer  and  occasional  clean-ups  of  cement  copper  made. 

Theoretically  but  0.88  Ib.  of  iron  is  required  to  replace 
one  pound  of  copper.  Generally,  however,  the  scrap 
contains  more  or  less  inert  graphite  and  iron  oxide,  the 
iron  precipitates  more  or  less  arsenic,  etc.,  and  the  high 
free  sulphuric  acid  actively  attacks  the  iron  so  that  as 
much  as  two  pounds  of  scrap  iron  are  often  required  per 
pound  of  copper  precipitate. 

The  cement  is  generally  quite  foul  running  about  as 
shown  in  Table  29: 

TABLE  29. — ANALYSIS  OF  TYPICAL  CEMENT  FROM  DISCARDED  ELECTROLYTE 


Copper, 
per  cent 

Iron 
per  cent 

Arsenic, 
per  cent 

Silver, 
oz.  per  ton 

Gold, 
oz.  per  ton 

70 

5 

10 

15 

0.1 

The  silver  and  gold  come,  of  course,  from  suspended 
slimes  carried  over  from  the  electrolytic  tanks. 

While  this  method  has  the  advantages  of  simplicity  in 
operation  and  small  plant  required,  it  makes  a  foul  cement 


92 


COPPER  REFINING 


which  requires  retreatment,  runs  up  a  heavy  bill  for  scrap 
iron  and  entirely  wastes  both  the  free  and  combined  sul- 
phuric acid  content  of  the  electrolyte. 

Nevertheless,  it  is  still  used  in  some  small  plants,  and  is 
always  considered  a  legitimate  emergency  measure  for 
dealing  with  a  bad  electrolyte. 

II.  The  By-product  Manufacture  of  Blue  stone. —The 
next  purification  method  developed  was  to  go  into  the 
manufacture  of  commercial  bluestone,  using  electrolyte  as  a 


Anode 

Electrolyte    Copper 
Shot 


Mother  Liquor 

r  *  •  > 

Shot  Towtr 

Scrap  Iron 

1 

h 

£ 
3 

•*\  Loilin*  Tank 

J 

T3 

02 

CrytzUlizing 
Tanks 

Mother      i  -"-  

Cement  Copper 


Bluestone     Fines  Waste  Liquor 

Crystals  Impurities 

FIG.  23. — Flow-sheet  of  by-product  bluestone  plant. 


raw  material.  The  process  consisted  of  four  steps,  as 
shown  in  Fig.  23:  Neutralization  of  free  acid  by  means  of 
anode  copper,  concentration  of  neutral  liquor  by  boiling, 
crystallization  of  heavy  liquor,  and  cementation  of  mother 
liquors. 

In  this  way  the  entire  free  acid  content  as  well  as  that 
already  combined  with  copper  in  the  electrolyte  is  recovered 
as  sulphate  of  copper  except  that  discarded  in  the  final 
mother  liquors  and  the  iron  tanks  do  not  receive  any 
liquor  until  the  impurities  have  risen  to  a  point  where  blue- 


IMPURITIES  93 

stone  of  commercial  purity  can  no  longer  be  made  by 
fractional  crystallization. 

Also,  as  anode  copper  is  used  for  neutralizing  the  free  acid 
a  certain  amount  of  copper  is  refined,  the  silver  and  gold 
remaining  as  slime  in  the  shot  towers,  and  therefore  the  by- 
product plant  is  entitled  to  a  certain  amount  of  tolls  to  be 
credited  against  its  operating  expense. 

Then  there  is  still  the  legitimate  profit  from  the  market 
price  of  bluetone,  so  that  it  is  possible  to  place  the  purifica- 
tion system  on  a  revenue-producing  basis. 

On  the  other  hand,  the  plant  required  is  bulky,  it  is 
dependent  for  results  upon  the  market  for  bluestone  which 
at  times  is  badly  overproduced  and  finally  the  plant  is 
strictly  limited  to  a  predetermined  capacity,  and  therefore 
uses  the  electrolyte  as  a  reservoir  for  fluctuating  amounts 
of  anode  impurities.  Also  where  very  foul  anodes  are  to 
be  treated  the  bluestone  plant  becomes  enormous. 

However,  as  this  system  is  more  or  less  used  to-day, 
though  generally  in  addition  to  other  methods,  some  ac- 
count of  the  process  will  be  given. 

1.  Capacity. — A  building  50  ft.  by  150  ft.  ground  plan 
will  produce  about  140,000  Ib.  of  bluestone  a  month.     This 
is  equivalent  to  about  35,000  Ib.  of  copper,  but  about  three- 
quarters  of  this  comes  from  the  shot  copper  and  only  9000 
Ib.  is  from  the  electrolyte,  so  that  this  plant  would  repre- 
sent the  withdrawal  of  but  100  cubic  feet  a  day  of  electrolyte. 

2.  Shot  Towers. — Copper  is  not  readily  soluble  in  dilute 
sulphuric  acid,  especially  in  the  presence  of  various  impuri- 
ties.    In  order  to  promote  the  rate  of  solution  the  liquor 
is  heated  and  poured  over  the  copper  intermittently  so  as  to 
promote  the  oxidation  of  the  copper.     The  towers  are 
built  in  various  ways  in  order  to  accomplish  this,  but  the 
simplest  method  is  to  spray  the  hot  liquor  over  the  top  of 
the  bed  and  let  it  trickle  through. 

The  copper  itself  is  cast  in  the  form  of  so-called  shot  in 
order  to  increase  the  surface  exposed.  Shot  is  a  mass  of 
small  hollow  irregular  spheres,  and  as  these  dissolve  away, 
the  mass  crumbles  and  this  tends  to  prevent  packing.  It  is 


94  COPPER  REFINING 

made  by  adding  a  little  matte  to  an  anode  furnace  charge 
to  lower  the  pitch  and  then  pouring  a  thin  stream  of  metal 
through  a  blast  of  air  into  a  well  filled  with  water.  More  or 
less  minor  explosions  occur,  but  if  the  procedure  is  skillfully 
carried  out  the  result  is  a  surprisingly  light  mass  of  detached 
globules  ranging  from  y±  in.  to  \Y%  in.  in  diameter. 

Underneath  the  shot  towers  is  a  well,  and  here  the  silver 
and  gold  slimes  wash  down  as  the  copper  is  dissolved  and 
settle  out.  These  are  periodically  collected  and  sent  to 
the  silver  refinery  to  be  smelted  along  with  the  anode  slimes 
from  the  electrolytic  tank  house.  It  is  impossible  ab- 
solutely to  neutralize  the  free  acid  in  the  electrolyte  in  this 
way  in  a  reasonable  time,  but  by  repeated  circulation  at 
high  temperature  it  is  possible  to  bring  it  below  1  per  cent. 
The  temperature  must  be  above  150°F.  for  effective  work, 
and  the  time  required  is  from  four  to  seven  days  in  a  plain 
tower. 

3.  Boiling  Tanks. — The  boiling  tanks  are  lead-lined  open 
tanks  with  closed  steam  coils  and  the  liquors  from  the  shot 
towers  are  here  concentrated  as  far  as  steam  at  a  reasonable 
pressure  for  lead  pipes  will  carry  them.     This  is  about  38° 
Beaume  and  corresponds  to  9  or  10  per  cent  copper.     The 
time  required  for  a  batch  is  about  eight  hours. 

4.  Crystallizing    Tanks. — The    crystallizing    tanks    are 
shallow,   open,   lead-lined   tanks  provided   with  strips  of 
lead  hung  from  cross  bars  on  which  the  crystals  grow.     The 
liquor  from  the  boiling  tanks  is  allowed  to  stand  quiescent  in 
these  tanks  for  five  days,  which  time  experience  has  shown 
to  be  the  best.     A  longer  period  results  in  more  or  less 
resolution,  while  forced  cooling  by  means  of  cooling  coils 
results  in  a  crop  of  small  and  less  pure  crystals. 

The  choice  large  crystals  grow  chiefly  as  trees  on  the  lead 
strips.  At  the  bottom  of  the  tanks  a  mass  of  fine  crystals 
forms. 

The  crystals  are  removed  after  the  mother  liquor  has 
been  pumped  out  and  are  then  generally  dried  by  hot  air 
and  screened  to  separate  the  coarse  and  fine,  each  class 


IMPURITIES  95 

being  packed  in  barrels  for  shipment.  The  fine  sand  is  sent 
back  to  the  boiling  tanks  and  recrystallized. 

The  best  crystals  physically  are  made  from  truly  neutral 
liquor.  Should  the  liquor  carry  2  per  cent  acid  or  over, 
the  crystals  are  likely  to  be  hygroscopic  and  are  not  a 
prime  market  product.  Also  calcium  sulphate,  if  present  in 
the  liquor  to  any  extent  (this  salt  is  sparsely  soluble  in 
such  solutions),  will  form  white  "  whiskers"  on  the  crystals. 

The  main  physical  question  is,  therefore,  to  obtain  a 
maximum  proportion  of  large,  clean,  dry  crystals.  An 
ordinary  yield  is  47  per  cent  large  crystals,  9  per  cent  pea 
and  44  per  cent  sand. 

Chemically  the  problem  is  to  keep  the  mother  liquor 
down  in  impurities  to  a  point  where  fractional  crystallization 
is  sharp  enough  to  make  a  commercial  salt.  The  alterna- 
tive is  to  dissolve  and  recrystallize.  The  effect  of  recrystal- 
lizing  is  shown  in  Table  30,  where  "bottoms"  have  been 
dissolved  and  crystallized,  forming  crystals  and  a  second 
crop  of  bottoms,  but  leaving  most  of  the  arsenic  and  anti- 
mony behind  in  the  new  mother  liquor. 

TABLE  30. — REMOVAL  OF  IMPURITIES  FROM  BLUESTONE  BY 
RECRYSTALLIZATION 


Per  cent 
arsenic 

Per  cent 
antimony 

Original  "bottoms" 

0  434 

0.113 

New  crystals  

0.076 

0.013 

New  bottoms  

0.117 

0.014 

CuSO4.5H2O  contains  approximately  25.4  per  cent  copper, 
38.5  per  cent  combined  acid  and  36.1  per  cent  water  of 
crystallization.  Ninety-nine  per  cent  crystal  should,  there- 
fore, run  slightly  better  than  25  per  cent  copper  and  98  per 
cent  slightly  less  and  bluestone  is  generally  considered  to 
carry  one-quarter  copper.  In  Table  31  are  given  some 
random  analyses  of  the  products  from  several  plants. 


96  COPPEP  REFINING 

TABLE  31. — REPRESENTATIVE  ANALYSES  OF  BLUESTONE 


Per 

Pei- 

Per 

Per 

Per 

Source 

Remarks 

cent 

cent 

cent 

cent 

cent 

copper 

iron 

nickel 

arsenic 

antimony 

A 

Coarse  

24.97 

0.195 

0.358 

0.018 

0.006 

A 

Fines  

24.72 

0.215 

0.419 

0.040 

0.009 

A 

Seconds 

23  25 

0  250 

1  250 

0  090 

0  017 

B 

"Anhydrous" 

26.18 

0.600 

1.850 

0.668 

0.030 

c 

Coarse 

25  22 

0  025 

0  150 

0  003 

0  005 

D 

Coarse  

24.80 

Trace 

0.150 

0.018 

The  wide  range  of  the  impurities  in  the  different  products 
in  Table  31  is  due  chiefly  to  the  difference  in  the  electrolytes 
at  the  various  plants.  The  salt  showing  26.18  per  cent 
copper  had  been  overdried,  thereby  losing  some  of  the 
water  of  crystallization;  this  makes  a  dirty  white  crystal 
and  is  also  a  source  of  loss  to  the  producer,  as  copper  con- 
tents above  the  guaranteed  amount  are  not  paid  for. 

5.  Mother  Liquors.- — The  mother  liquor  from  the  crystal- 
lizing tanks  are  returned  to  the  boiling  tanks  and  reconcen- 
trated  with  fresh  liquor  from  the  shot  towers.  As  the 
impurities  grow  in  concentration,  however,  the  bluestone 
is  increasingly  contaminated  and  a  point  is  reached  where  a 
proportion  has  to  be  diverted  to  the  iron  tanks  to  be  worked 
up  into  cement  copper. 

Sometimes  a  foul  copper-nickel-iron  sulphate  is  crystal- 
lized from  such  liquors  before  cementing.  An  example  of 
such  a  salt  is  shown  in  Table  32. 

TABLE   32. — COPPER,    NICKEL,    IRON    SULPHATE   FROM    MOTHER   LIQUOR 


Per  cent 
copper 

Per  cent 
nickel 

Per  cent 
iron 

Per  cent 
arsenic 

Per  cent 
antimony 

Oz.  per  ton 
silver 

9.52 

11.50 

1.19 

0.23 

0  .  024 

0.15 

The  silver  is,  of  course,  due  to  contamination  by  sus- 
pended slimes  from  the  shot  towers.  Similar  quantities 
and  a  proportionate  amount  of  gold  is  found  in  bluestone 


IMPURITIES  97 

when  anode  copper  is  used  for  shot.  A  salt  of  this  charac- 
ter has  no  commercial  use,  but  is  a  good  starting  point  for 
nickel  recovery  as  another  by-product. 

III.  Purifying  With  Insoluble  Anode  Tanks. — Where  but 
little  nickel  is  present  and  arsenic  is  the  predominant  im- 
purity, a  simple  process  based  upon  the  use  of  insoluble 
anode  tanks  is  sometimes  employed. 

A  certain  amount  of  electrolyte  is  diverted  to  a  set  of 
tanks  provided  with  lead  anodes  and  operating  in  cascade, 
so  that  practically  all  of  the  copper  and  arsenic  are  plated 
out  and  the  equivalent  sulphuric  acid  is  liberated.  The 
treated  liquor  is  returned  to  the  electrolyte. 

This  uses  a  considerable  amount  of  power,  but  recovers 
the  acid  and  calls  for  no  scrap  iron.  It  takes  three  tanks 
in  cascade  to  bring  the  copper  and  arsenic  down  to  0.1 
per  cent  or  less.  The  first  tank  will  operate  at  about  85 
per  cent  current  efficiency  and  produce  cathodes,  which 
may  with  reasonable  safety  be  included  in  the  fine  copper 
output ;  the  second  will  run  at  perhaps  50  per  cent  efficiency 
and  the  cathodes  reserved  for  casting  copper  or  anode 
changes;  the  last  tank  runs  at  a  very  low  copper  efficiency 
and  produces  a  sludge  consisting  of  about  half  copper  and 
half  metallic  arsenic.  Much  of  the  latter  can  be  eliminated 
from  this  by  roasting  and  sublimation,  if  desired.  Such 
insoluble  anode  tanks  evolve  arseniuretted  hydrogen,  which 
is  very  poisonous,  and  they  should  therefore  be  located  out 
of  doors. 

This  method  is  limited  by  the  permissible  amount  of 
impoverishment  of  the  electrolyte  in  copper;  if  carried 
to  extremes  it  is  evident  that  the  whole  copper  content  of 
the  electrolyte  would  be  withdrawn,  the  acid  being  corre- 
spondingly increased.  On  the  other  hand,  when  nickel 
is  practically  absent  from  the  anodes  and  the  copper 
withdrawals  necessary  to  hold  down  the  arsenic  do  not 
exceed  the  normal  growth  of  copper  in  the  electrolyte,  it  is  a 
satisfactory  solution  of  the  purification  problem. 

IV.  Complete    Cyclical    Purification. — The    method    of 
purification  of  the  electrolyte  in  general  use  among  copper 


98 


COPPER  REFINING 


refineries  to-day  is  shown  in  diagrammatic  form  in  Fig. 
24.  Electrolyte  is  diverted  to  insoluble  anode  tanks  at  a 
rate  just  sufficient  to  keep  the  determining  impurity  at  the 
desired  point.  This  impurity  is  almost  always  nickel 
or  arsenic  and  generally  the  former. 


FIG.  24. — Modern  method  of  purifying  electrolyte. 


The  first  step  in  the  process  is  the  same,  therefore,  as 
that  just  described  in  the  preceding  paragraph,  but  the 
liquor  resulting  therefrom  instead  of  being  returned  to  the 
electrolyte  is  sent  to  a  steam  boiling  tank  where  it  is  con- 
centrated as  in  the  manufacture  of  bluestone  up  to  about 
40°  Beauine*. 

The  liquor  is  then  transferred  to  a  tank  made  of  boiler 
plate,  as,  being  nearly  copper-free  and  of  high  sulphuric 
acid  content,  it  is  no  longer  seriously  corrosive  to  iron. 
This  evaporator  is  heated  by  direct  fire  until  the  syrup 


IMPURITIES  99 

reaches  about  70°  Beaume,  at  which  strength  practically 
all  the  impurities  except  the  small  amount  of  arsenic  which 
has  escaped  the  insoluble  anode  tanks  and  the  sodium  and 
potassium  salts  have  been  precipitated  as  anhydrous 
sulphates. 

This  heavy  liquor  with  its  suspended  solids  is  then 
tapped  on  to  a  sand  filter  where  the  bulk  of  the  strong 
acid  is  filtered  out.  The  mushy  salts  are  then  shoveled 
on  to  a  draining  board  and  finally  into  a  sucking  tub, 
where  the  acid  is  displaced  with  a  very  small  quantity 
of  w^ater. 

The  partly  washed  salts  are  then  shoveled  on  to  a  drying 
floor  where  the  water  is  gradually  taken  up  as  water  of 
crystallization  and  the  mass  sets  into  lumps  of  partially 
hydrated  sulphate,  which  may  be  readily  handled  and 
shipped  and  is  in  good  physical  shape  for  charging  into  a 
furnace  for  the  recovery  of  metal  values. 

If  the  strong  acid  filtrate  is  chilled  before  returning 
it  to  the  electrolyte  much  of  the  sodium  sulphate  will  be 
thrown  out.  Practically  this  occurs  at  ordinary  winter 
temperatures  at  most  of  the  plants. 

This  process  is  completely  cyclical  except  for  the  com- 
bined acid  sent  out  with  the  crude  nickel  salts  and  such 
acid  losses  as  may  be  incurred  by  fumes  from  the  fire 
evaporator.  The  nickel  is  recovered  in  a  form  reasonably 
acceptable  to  a  nickel  smelter;  but  little  low-grade  cathode 
copper  is  made,  and  the  arsenic  could  be  separately  recov- 
ered were  it  worth  while  to  do  so. 

The  objection  still  applies  that  the  electrolyte  may  be 
robbed  of  too  much  copper  and  this  has  especial  force  when 
discussing  nickel,  as  much  of  the  anode  nickel  appears  to 
dissolve  electrochemically  so  that  more  copper  is  plated 
out  at  the  cathode  than  is  electrochemically  dissolved  at 
the  anode. 

The  remedy,  in  case  of  trouble  of  this  character,  is  to 
build  shot  towers  and  allow  a  certain  proportion  of  the 
regular  electrolyte  to  trickle  through  them.  As  the  solu- 
tion so  diverted  has  always  high  free  acid  content,  such 


100 


COPPER  REFINING 


towers  are  more  active  than  in  a  bluestone  plant,  where 
complete  neutralization  is  the  object. 

The  chemical  separations  are  not  sharp  and  some  leeway 
has  to  be  left  for  circulating  impurities  due  to  this  fact. 
An  example  of  crude  nickel  salt  made  in  this  way  is  given 
in  Table  33. 

TABLE  33. — ANALYSIS  OF  CRUDE  NICKEL  SALT 


Per  cent 
water 

Per  cent 
copper 

Per  cent 
iron 

Per  cent 
nickel 

Per  cent 
arsenic 

17.0 

0.57 

1.76 

28.45 

0.02 

F.   RECOVERY   OF  INSOLUBLE  IMPURITIES    FROM  THE  ANODE 

SLIMES 

This  question  involves  the  whole  metallurgy  of  the  silver 
refinery  and  will  be  taken  up  separately  in  the  next  chapter. 


CHAPTER  VII 
BY-PRODUCTS 

In  the  last  chapter  the  various  elements  commonly  asso- 
ciated with  copper  were  discussed  as  impurities  and  their 
treatment  carried  far  enough  to  get  them  out  of  the  refined 
copper.  We  therefore  left  them  as  anode  slimes  and  puri- 
fying plant  salts  or  as  furnace  slags  and  flue  dust.  We 
have  now  to  discuss  the  working  up  of  these  impure  metal- 
lurgical products  into  marketable  materials. 

Market. — We  may  broadly  divide  the  possible  by-pro- 
ducts into  three  main  classes  as  to  their  probable  disposition 
as:  First,  those  which  are  readily  marketable  in  any  quan- 
tity which  such  sources  are  ever  likely  to  supply;  second, 
those  which  have  a  limited  and  therefore  sensitive  and 
widely  fluctuating  market,  and,  third,  those  which  have 
at  the  present  time  practically  no  market  at  all.  These 
groups  consist  of  the  following:  first,  gold,  silver,  platinum, 
palladium,  lead,  nickel  and  antimony;  second,  bismuth, 
arsenic,  cobalt,  selenium;  and,  third,  tellurium. 

Palladium  is  the  only  uncommon  member  of  class  one. 
The  production  of  this  element  is  as  yet  quite  small  and 
this  amount  is  readily  absorbed  by  the  jewelry  and  scien- 
tific-instrument trade.  Practically  it  has  the  useful  qualities 
of  platinum,  but  is  of  about  half  the  density. 

Bismuth  goes  chiefly  into  the  drug  trade.  Any  large 
production  of  it  floods  the  market  and  depresses  the  price. 
It  is  seldom  associated  with  copper  in  any  quantity,  al- 
though it  is  the  mainstay  of  electrolytic  lead  refining. 

There  is  a  large  market  for  arsenic,  which  is  used  chiefly 
for  its  poisonous  properties  in  insecticides,  etc.,  but  in  re- 
cent years  the  great  output  of  arsenious  oxide  recovered 
from  smelter  flue  dusts  has  glutted  this. 

101 


102  COPPER  REFINING 

Cobalt  is  an  element  that  has  not  yet  found  itself.  The 
Canadian  Government  has  conducted  a  wide  investigation 
as  to  its  possible  uses,  but  aside  from  the  use  of  cobalt 
oxide  in  the  manufacture  of  smalt  and  for  whitening  yellow 
pottery  and  the  use  of  the  metal  to  give  certain  minor 
properties  to  cobalt-plated  ware  and  cobalt  alloys,  no 
marked  characteristic  giving  it  a  superior  value  to  nickel 
has  been  discovered,  and  so  far  it  has  not  been  found  pos- 
sible to  produce  it  at  the  price  of  nickel.  Except  in  one  or 
two  notable  cases,  as  in  Katanga  bullion,  cobalt  is  not  asso- 
ciated with  copper  to  any  extent. 

Selenium  was  a  rare  element  a  few  years  ago ;  to-day  a  single 
copper  refinery  could  easily  produce  ten  tons  a  month, 
could  it  be  marketed.  Its  peculiar  photo-electric  property 
has  not  created  any  tonnage  demand,  and  its  only  major  use 
is  for  coloring  ruby  glass.  This  creates  a  steady  but  very 
moderate  market. 

Tellurium  could  also  be  produced  by  the  ton,  but  there 
appears  to  be  no  reasonably  extensive  market  for  it.  The 
various  plants  producing  selenium  and  tellurium  are  now 
joining  trying  to  develop  wider  markets  for  these  elements 
by  scientific  research.1 

Market  Requirements. — Gold  is  nearly  always  sold  to 
the  United  States  Assay  Office.  There  are  no  requirements 
as  to  fineness,  but  the  bullion  is  weighed,  sampled  and  as- 
sayed by  the  Government.  It  is  then  purchased  with  a 
"gold  check"  on  the  Treasury  Department  for  full  value 
less  certain  refining  charges.  These  charges  are  so  adjusted 
that  they  are  greater  than  those  of  an  outside  refiner,  unless 
the  gold  is  nearly  pure;  but  the  last  traces  of  impurities 
are  taken  out  by  the  government  for  less  than  it  would  cost 
others  who  are  properly  equipped.  This  is  due  to  the  fact 
that  the  government  is  not  concerned  with  interest  charges 
while  the  gold  is  in  process,  while  the  individual  is  very 
much  concerned  therewith.  The  result  is  that  it  is  cus- 
tomary for  the  refiner  to  bring  his  gold  up  to  perhaps  985 
parts  fine,  the  balance  being  chiefly  silver  (which  is  not 

1  See  Lenher,  Chem.  &  Met.  Eng..  vol.  22,  p.  1108. 


BY-PRODUCTS  103 

paid  for  above  992)  and  sometimes  small  quantities  of  the 
platinum  group  of  metals  which  are  likewise  forfeited. 

Silver  is  at  times  exported  to  London  when  the  English 
market  is  higher  than  the  American  by  an  amount  sufficient 
to  pay  transfer  charges.  America  also  makes  direct  ship- 
ments westward  to  the  market  in  the  Orient. 

The  silver  used  locally  goes  to  brokers,  but  little  of  the 
home  consumption  passing  directly  into  the  hands  of  the 
user. 

The  chief  impurity  in  bar-silver  is  copper.  For  the 
American  market  it  is  customary  t'o  produce  999  fine,  and 
for  the  English  998.  As  value  over  998  is  neglected  in  pay- 
ments from  abroad  while  penalties  are  exacted  if  it  runs 
under  that  point,  care  must  be  taken  to  bring  high  grade 
electrolytic  silver  down  in  fineness.  This  is  done  when  melt- 
ing by  the  addition  of  copper  wire  in  sufficient  quantity  to 
allow  a  safe  margin  above  998  for  assay  variations.  Also, 
as  only  even  quarter  ounces  are  paid  for  it  is  customary 
to  plug  the  bars  with  silver  to  a  safe  weight  above  an  even 
quarter.  The  same  result  could  be  obtained  by  trimming 
but  this  would  obscure  any  evidence  of  theft  by  trimming 
in  transit. 

Tellurium  is  the  mam  metallurgical  enemy  of  silver  and 
must  be  thoroughly  eKminated  from  it  in  order  to  prevent 
brittleness  in  subsequent  rolling  of  the  metal. 

Printed  schedules  of  the  charges  of  the  U.  S.  Mints  and 
Assay  Offices  are  periodically  issued  by  the  Treasury  Depart- 
ment and  may  be  obtained  upon  application.  The  points 
of  chief  interest  to  a  copper  refiner  in  the  schedule  in  effect 
July  1,  1919,  are:  first,  all  deposits  are  subject  to  a  melting 
charge  of  $1  per  1000  oz.;  second,  ordinary  dore  is  charged 
3-2  cent  an  ounce  for  parting  and  refining;  third,  gold  bul- 
lion between  950  and  991.75  parts  fine  is  charged  2  cents 
an  ounce  for  refining  while  silver  contents  in  such  bullion 
above  992  fine  are  not  paid  for;  fourth,  metals  of  the  plati- 
num group  are  not  paid  for,  although  then-  presence  in 
quantity  will  be  reported  for  the  information  of  the  deposi- 
tor, and  fifth,  government  weights  and  assays  control,  the 


104  COPPER  REFINING 

weights  being  to  one  hundredth  of  an  ounce  and  the  fine- 
ness to  the  nearest  quarter-part. 

The  platinum  group  of  metals  are  sold  as  a  rule  in  .crude 
shape  as  a  sponge  or  powder,  running  950  fine  or  better. 
In  passing,  it  may  be  of  interest  to  state  that  the  ordinary 
assay  methods  for  the  platinum  group  give  very  inaccurate 
results  and  that  a  true  assay  for  platinum  and  palladium 
is  a  tedious  and  costly  operation,  involving  repeated  pre- 
cipitations*in  order  to  free  the  platinum  metals  from  the  last 
traces  of  lead  sulphate  and  silver  chloride. 

Lead  is  seldom  present  in  any  quantity  unless  it  is  added 
in  the  cupeling  operation  in  the  silver  refinery.  Its  market 
requirements  are  too  well  known  to  need  discussion  here. 

Nickel  is  used  first,  as  nickel  sulfate  or  as  a  double 
ammonium  sulfate  in  nickel  plating;  second,  as  anodes; 
and  third,  as  shot  or  oxide  for  steel  alloying.  The  crude 
sulphate  of  nickel,  iron  and  copper  produced  by  a  purifying 
plant  can  be  made  into  a  reasonably  pure  sulphate  by  wet 
methods.  This  is  then  broken  down  by  heat  into  oxide 
which  is  in  turn  reduced  by  carbon  to  metallic  nickel  in  an 
electric  crucible.  Carbon-free  nickel  can  be  made  elec- 
trolytically  from  this  product  if  desired.  Nickel  anodes 
are  generally  alloyed  with  more  or  less  iron,  although 
conditions  in  the  trade  are  improving  in  this  respect  with  a 
corresponding  improvement  in  the  character  of  nickel 
plating.  The  production  of  pure  nickel  salts  is  very 
troublesome  and  should  not  be  attempted  unless  unusually 
pure  electrolyte  is  available  as  a  starting  point.  One  of  the 
large  manufacturers  of  this  product  had  so  much  trouble 
in  meeting  the  strict  requirements  of  the  trade  that  wet 
methods  of  production  were  abandoned  and  the  salt  pro- 
duced by  dissolving  pure  nickel. 

Electrolytic  nickel  should  run  99.80  nickel  plus  cobalt 
and  should  be  free  from  carbon  and  oxides.  The  cobalt 
should  not  exceed  0.3  per  cent  and  the  sulphur  should  be 
very  low.  Small  quantities  of  iron  and  copper  are  per- 
missible and  a  little  arsenic  is  generally  present. 

Table  34  gives  some  examples  of  the  crude  by-product 


BY-PRODUCTS 


105 


salt  as  made  by  the  refinery  which  is  the  raw  material 
for  by-product  nickel;  also  analyses  of  commercial  single 
and  double  salts  as  found  in  the  market. 

TABLE  34. — ANALYSES  OF  NICKEL  SALTS 


Per 

Per 

Per 

Per 

Pei- 

Per 

Per 

Per 

Source 

Type 

cent 
Ni 

cent 
Cu 

cent 
Zn 

cent 
Fe 

cent 
Co 

cent 

As 

cent 
Sb 

cent 
Pb 

A 

Crude 

25.32 

0.82 

0.19 

3.10 

1.27 

0.29 

0.05 

B 
B 

Single 
Double 

22.34 
14.93 

None 
0.008 

None 
None 

0.002 
Trace 

0.016 
None 

None 
None 

None 
None 

C 
C 

Single 
Double 

20.80 
14.80 

Trace 
None 

0.001 
None 

0.007 
0.003 

0.053 
0.067 

None 
None 

None  ' 
None 

None 
None 

An  example  of  nickel  made  in  the  electric  furnace  from 
refinery  by-product  salt  showed  98.98  per  cent  nickel, 
0.29  per  cent  copper  and  0.73  per  cent  iron. 

Antimony  is  marketed  chiefly  in  " metallic"  form  as  its 
principal  use  is  in  the  manufacture  of  alloys.  It  should  be 
sufficiently  pure  to  show  clearly  marked  "  stars"  on  the  sur- 
face of  the  pigs.  It  can  also  be  marketed  as  a  "  hard"  lead. 

Bismuth  is  produced  as  metal  in  very  pure  form,  usually 
obtained  by  electrolytic  refining. 

Arsenic  is  marketed  as  a  99  per  cent  arsenious  oxide 
readily  obtained  by  sublimation  from  many  flue  dusts. 

Cobalt  is  marketed  almost  entirely  as  oxide.  There 
are  three  oxides  of  cobalt  and  all  have  a  market  demand. 
They  are  difficult  to  prepare  and  the  trade  requirements 
are  based  on  previous  production  rather  than  on  any 
rational  specification.  The  great  use  of  cobalt  oxide  is 
as  an  offset  to  yellow  cast  in  clay  and  similar  material, 
the  powerful  blue  characteristic  of  cobalt  salts  being  equiva- 
lent to  a  bleach. 

Selenium  is  produced  in  granular  or  fused  form,  99 
per  cent  pure.  The  black  variety,  not  too  finely  ground, 
f  s  desired  by  the  glass  industry. 


106 


COPPER  REFINING 


Tellurium  is  usually  offered  in  an  impure  state  but  no 
market  standards  have  as  yet  been  developed. 

Metallurgy. — The  general  treatment  of  the  electrolyte 
for  the  recovery  of  a  crude  nickel  sulphate  has  already  been 
discussed  in  Chapter  VI.  This  method  also  gave  a  cathode 
sludge  of  arsenic  and  copper  from  which  a  certain  amount  of 
arsenious  oxide  could  be  recovered  if  desirable.  The 
balance  of  the  impurities  were  either  lost  in  blast  furnace 
slags  or  concentrated  in  anode  slimes. 

TABLE  35. — ANALYSES  OF  ANODE  SLIMES 


Per  cent 

Raw 

Boiled 

A 

B* 

C 

D 

E 

Copper.  . 

14.3 
35.0 
0.64 

43.3 
17.2 
0.12 
0.00017 

20.0 
37.0 
0.6 

1.60 
41.5 
0.7 
0.0007 
0.0006 
0.89 

1.14 
29.5 
0.7 

Silver. 

Gold  
Platinum  
Palladium 

1.07 

Nickel  

5.25 

0.08 
0.006 
3.03 
3.46 
0.11 
13.21 
0.36 
0.76 
1.20 
2.10 
0.09 
0.18 

Cobalt.. 

Arsenic  

2.68 
5.35 
0.46 
1.69 
0.17 
2.44 
5.70 
2.69 
Trace 
4.40 

4.0 

8.0 

1.42 
3.84 
0.37 
2.48 
0.24 
7.33 
12.94 
5.72 
Trace 
5.29 

1.20 
5.70 
0.20 
1.97 
0.26 
18.60 
11.24 
6.20 
Trace 
4.71 

Antimony  

Bismuth  
Sulphur  .  .           .... 

Iron  

Lead  . 

Selenium  
Tellurium   . 

Zinc. 

Silica  

The  analyses  given  in  Table  35  are  representative  of 
such  slimes.  The  wide  variation  in  the  composition  is 
due  to  different  conditions  obtaining  at  different  plants  and 
this  lack  of  a  uniform  product  has  been  one  of  the  diffi- 
culties in  the  way  of  a  wet  treatment  of  slimes.  A  is  an 
ordinary  slime  with  nothing  unusual  in  the  composition. 
B  is  produced  by  the  electrolysis  of  converter  anodes; 
its  nature  is  reflected  in  the  high  copper  and  sulphur  con-* 


BY-PRODUCTS  107 

tents.  D  and  E  are  analyses  after  the  slime  has  been 
somewhat  oxidized  and  leached  with  dilute  sulphuric 
acid  to  remove  the  excess  copper;  they  show  noteworthy 
amounts  of  selenium  present.  The  high  lead  content  of  E 
results  from  anodes  carrying  0.2  per  cent  lead. 

There  is  room  for  division  of  opinion  as  to  the  exact 
compounds  present  in  anode  slimes.  The  silver,  gold,  and 
a  large  part  of  the  copper  in  slimes  from  well-refined  anodes 
are  doubtless  in  metallic  or  alloy  form.  Selenium  and 
tellurium  may  be  present  in  elemental  form  or  as  selenides, 
etc.  Lead  is  probably  as  sulphate  and  much  of  the  nickel, 
iron,  cobalt,  arsenic,  antimony  and  bismuth  in  some 
oxidized  form,  but  antimoniates  and  analogous  compounds 
as  well  as  simple  oxides  may  easily  be  present.  In  general 
the  slime  is  in  an  oxidized  condition  since  the  prior  proc- 
esses in  the  converter,  the  anode  furnace  and  during 
electrolysis  are  all  strongly  oxidizing.  The  less  noble 
metals  are  all  therefore  more  or  less  oxidized  in  the  slime 
when  drawn  from  the  electrolytic  tanks. 

There  is  always  a  high  percentage  of  copper  present. 
In  slimes  from  converter  anodes  this  is  due  in  part  to 
undecomposed  matte,  but  even  in  slimes  from  the  purest 
anodes  there  is  a  considerable  amount  of  very  finely  divided 
metallic  copper  to  be  found  after  all  cathode  needles,  etc., 
have  been  carefully  screened  out.  This  probably  comes 
from  either  mechanical  disintegration  of  the  anode  under 
electrolysis  or  from  chemical  oxidation  of  cuprous  salts 
formed  at  the  anode.  Any  one  interested  in  the  various 
possible  explanations  of  this  action  would  do  well  to  consult 
an  article  by  Dr.  Emil  Wohlwill  on  anode  disintegration 
in  the  1903  volume  of  Zeitschrift  fur  Electrochemie. 

Requirements  for  a  Slimes  Process. — Perhaps  the 
foremost  requirement  for  a  metallurgical  process  for  the 
treatment  of  these  complex  slimes  is  that  there  shall  be  a 
minimum  of  loss  of  the  precious  metals.  We  are  dealing 
with  material  which  usually  runs  35  or  40  per  cent  silver 
and  half  a  per  cent  of  gold  (and  at  times  much  more).  This 
means  that  every  by-product  must  be  retreated  for  values 


108  COPPER  REFINING 

contained.  The  second  question  relates  to  the  interest  on 
metal  value.  The  slimes  are  worth  about  $5  a  pound  and  a 
process  is  needed  which  will  turn  out  the  great  bulk  of  the 
gold  and  silver  promptly.  A  third  desideratum  is  to  be 
able  to  recover  as  by-products  any  marketable  impurities 
which  are  present  in  commercial  amounts.  Then  we 
have  to  consider  the  chemical  questions  introduced  in  the 
parent  process  of  copper  refining  by  the  presence  of  any 
unusual  reagents.  Finally  we  have  to  reckon  with  the 
operating  cost  of  the  process  in  question. 

This  problem  of  treating  anode  slimes  is  doubtless  the 
most  difficult  presented  by  the  metallurgy  of  copper  refin- 
ing, and  it  is  also  the  one  which  to-day  leaves  most  room  for 
further  improvement.  It  has  always  afforded  a  field  of 
contest  for  advocates  of  wet  against  advocates  of  fire  pro- 
cesses; and  the  methods  used  at  present  employ  a  mixture 
of  the  two,  the  wet  processes,  however,  steadily  encroach- 
ing as  time  passes.  We  shall  first  discuss  present  practice 
and  then  the  possible  points  of  attack  for  further  improve- 
ment. 

The  general  problem  as  at  present  handled  involves 
three  main  steps:  first,  the  production  of  a  nearly  copper- 
free  slime;  second,  smelting  this  to  dore  bullion;  and  third, 
parting  the  dore. 

Slimes  Boiling. — As  sulphuric  acid  is  a  cheap  reagent  and 
as  sulphate  of  copper  is  the  basis  of  the  electrolyte,  the 
obvious  method  by  which  to  remove  the  copper  from  the 
anode  slime  is  to  convert  it  into  sulphate.  As  copper  is  not 
readily  soluble  in  dilute  sulphuric  acid,  however,  either 
strong  acid  or  some  oxidizing  agent  such  as  air  or  nitre 
must  be  used. 

The  practice  in  the  earlier  plants  was  to  screen  the  slimes 
through  60  or  70  mesh  copper  wire  into  agitators,  using 
water  to  force  the  slimes  through  the  screen.  These  agita- 
tors were  lead-lined  tanks  about  six  feet  in  diameter  and 
six  feet  deep,  carrying  a  two-bladed  paddle  2^  in.  from  the 
bottom  and  driven  12  rotations  per  minute.  The  blades 
tended  to  raise  the  slimes  from  the  bottom.  A  charge 


BY-PRODUCTS  109 

consisted  of  perhaps  1000  Ib.  of  slimes.  After  these  were 
sluiced  down  into  the  tank  liquor  was  siphoned  off  and 
enough  sulphuric  acid  added  to  make  a  50  per  cent  solution. 
A  steam  pipe  heated  the  charge  to  the  boiling  point,  the 
paddles  were  started  and  100  Ib.  soda  nitre  added,  a  few 
pounds  at  a  time.  Owing  to  the  strong  evolution  of  poison- 
ous nitrous  fumes  the  nitre  was  added  only  at  night  and  the 
treatment  extended  over  two  and  sometimes  three  nights 
when  the  copper  was  found  to  be  reduced  to  perhaps  3 
per  cent.  The  liquors  made  were  sent  to  the  electrolyte 
after  carefully  settling  to  free  them  from  slimes. 

After  agitating,  washing,  and  settling  the  charge,, the 
supernatant  liquor  was  siphoned  off  and  the  slimes,  now 
running  about  60  per  cent  moisture,  dried  to  a  thick  mud 
carrying  25  per  cent  moisture  by  means  of  closed  steam 
coils.  This  mud  was  considered  ready  for  charging  into 
the  reverberatory  refining  furnace. 

This  method  has  been  subject  to  many  variations  in 
details.  At  some  plants  compressed  air  is  used  for  agita- 
tion instead  of  the  paddles.  Drying  has  been  accomplished 
by  the  use  of  filter  presses,  of  centrifugals  or  by  direct  heat 
instead  of  steam.  It  remained  the  general  method,  how- 
ever, until  the  development  of  cyclical  purification  of  the 
electrolyte  which  system  made  objectionable  the  introduc- 
tion of  large  quantities  of  sodium  salts  into  the  electrolyte. 

This  difficulty  was  first  met  by  charging  raw  slimes  high 
in  copper  directly  into  the  dore  furnace.  As  will  be  later 
explained,  such  a  procedure  locks  up  a  large  quantity  of 
silver  in  circulating  by-products  and  is  very  undesirable. 

The  next  step  was  the  development  of  slimes-roasting  and 
air  has  now  come  into  general  use  as  an  oxidizing  agent. 
In  some  cases  an  ordinary  roasting  hearth  is  used  and  in 
others  patented1  forms  of  tray  hearths,  but  in  any  case  all 
that  is  necessary  is  to  expose  a  shallow  layer  of  slimes  to  a 
low  heat  with  sufficient  air  passing  to  hold  the  temperature 
below  the  fritting  point  (the  slimes  often  ignite  and  over- 
heat from  the  heat  of  oxidation),  while  being  very  careful 

1  Wales,  U.  S.  Patent. 


110  COPPER  REFINING 

not  to  stir  the  bed,  and  produce  dust  losses.  The  resulting 
calcines  are  then  leached  with  dilute  sulphuric  acid  which 
readily  reacts  with  the  cupric  oxide  formed  in  the  roasting. 

A  third  process,  which  is  patented,1  mixes  the  filter- 
pressed  slimes  with  concentrated  sulphuric  acid  in  sufficient 
quantity  to  satisfy  the  copper  present,  and  heats  this  mud 
in  a  furnace  the  hearth  of  which  is  a  steel  basin.  The  slimes 
and  acid  are  mixed  first  and  pumped  into  the  furnace. 
The  mass  is  heated  to  about  450°F.  with  an  oil  burner  and 
the  mud  occasionally  stirred.  The  residues  are  practically 
dry  and  yield  their  copper  sulphate  readily  when  digested 
in  hot  water.  This  operation  uses  but  a  chemical  equiva- 
lent of  acid  which  is  an  advantage  in  some  cases. 

These  various  oxidizing  and  leaching  processes  designed 
to  separate  the  copper  from  the  rest  of  the  slimes  can  be 
made  to  bring  the  residues  down  to  1  per  cent  or  less  in 
copper,  with  careful  handling.  The  other  elements  present 
are  also  more  or  less  affected  by  the  operation.  During 
the  process  more  or  less  silver  sulphate  is  formed,  but  this  at 
once  reacts  with  the  copper  present  to  form  copper  sul- 
phate and  metallic  silver.  In  fact  boiling  is  generally 
continued  until  the  presence  of  silver  in  solution  is  shown 
by  a  salt  test.  When  a  decided  precipitate  is  shown  a  small 
quantity  of  raw  slimes  is  added  to  precipitate  this  silver 
and  it  is  then  certain  that  sufficient  work  has  been  done  to 
oxidize  practically  all  of  the  copper.  In  roasting  the  more 
volatile  elements  sublime  to  a  certain  extent  and  such  flue 
dusts  are  a  fruitful  source  of  selenium,  for  example.  Solu- 
tions from  the  boiling  tanks  always  show  nickel,  arsenic  and 
antimony  extraction. 

Smelting  to  Dore. — Whatever  be  the  method  employed 
in  leaching  the  bulk  of  the  copper  from  the  slimes,  the 
residue  is  always  smelted  to  dore  in  some  type  of  reverbera- 
tory  furnace.  Were  the  slimes  subjected  to  a  quiet  fusion 
three  layers  of  molten  material  would  form  with  a  distinct 
separation  by  specific  gravity.  At  the  bottom  would  be  a 
foul  dore  bullion;  in  the  middle  would  be  what  could  be 

1  Keller,  U.  S.  Pat.  No.  1110493. 


BY-PRODUCTS 


111 


called  a  matte,  silver  being  the  principal  metallic  component 
and  selenium  substituting  for  the  usual  sulphur;  then  on 
top  would  be  an  arsenic-antimony  slag. 

It  is  quite  possible  to  base  a  process  on  such  a  fusion  of 
raw  slimes  with  subsequent  separation  and  special  treat- 
ment for  the  three  resultant  products.  However,  we  should 
get  but  about  half  of  the  silver  and  most  of  the  gold  hi 
a  bullion  around  900  fine,  while  most  of  the  balance  of  the 
silver  would  be  tied  up  in  the  copper-silver-selenium-tel- 
lurium matte.  This  matte  is  a  troublesome  material  to 
handle  without  making  undue  losses  or  tie-up  of  silver,  and 
it  is  because  the  quantity  made  is  more  or  less  proportional 
to  the  amount  of  copper  present  that  it  is  desirable  to 
eliminate  this  last  element  as  far  as  possible  before  smelting. 

The  earlier  work  was  done  in  cupels  with  the  addition  of 
lead,  the  process  being  in  a  cycle  of  charging,  melting,  skim- 
ming, cupeling  with  a  blast,  skimming  and  pouring.  Some- 
thing like  10  per  cent  of  the  lead  used  in  this  process  would 
be  lost,  and  the  resulting  slag  (Table  36)  would  have  to  be 
retreated  on  a  lead  basis  and  therefore  was  generally  shipped 
elsewhere. 

TABLE  36. — ANALYSIS  OF  LEAD  CUPEL  SLAG 


Gold 

Silver 

Copper 

Lead 

Iron 

Silica 

Skimming  

0.0017 

1.62 

5.42 

32.51 

4.87 

14.02 

Then  the  use  of  lead  was  abandoned  in  favor  of  straight 
reverberatory  smelting  in  small  basic  furnaces,  air  being 
forced  under  the  surface  of  the  bath  to  promote  oxidation. 
It  is  interesting  to  note  in  passing  that  the  silver  furnaces 
of  to-day  are  as  large  as  the  copper  reverberatories  in  the 
earliest  refineries.  The  general  plan  of  operations  is  to 
melt  the  charge,  thin  the  first  slag  with  such  fluxes  as  silica, 
soda  ash,  salt  cake,  etc.,  skim,  blow  the  small  quantity  of 
matte  to  a  second  slag  and  bullion,  skim,  fine  the  dore  with 
air  and  nitre,  skim,  and  pour.  Table  37  gives  representative 


112  COPPER  REFINING 

TABLE  37. — ANALYSES  OF  DORE  FURNACE  PRODUCTS 


Per  cent 

Slimes 

1st  slag 

2d  slag 

Dore 

Flue 
dust 

Copper  

2  39 

4  25 

7  90 

1  34 

0  1 

Silver.... 

40  0 

4  03 

3  31 

96  345 

5  0 

Gold 

0  767 

0  041 

0  0041 

2  37 

0  017 

Nickel  

1  94 

7  58 

1  02 

0  0065 

Arsenic  .... 

2  34 

1  593 

0  73 

0  0075 

2  65 

Antimony 

5  91 

11  34 

3  30 

0  0084 

10  9 

Bismuth  
Sulphur 

0.343 
2  80 

0.5378 
0  264 

0.64 
0  90 

Trace 

Iron.  . 

0  23 

5  63 

3  88 

0  1233 

Lead  
Selenium 

5.43 
11  89 

6.14 
1  58 

2.78 
11  06 

0  0020 

35 

Tellurium 

5  45 

1  80 

19  36 

0  0068 

1  5 

Zinc  

Trace 

Trace 

Trace 

None 

Silica  . 

4  25 

19  90 

2  26 

Alumina  
Magnesia  . 

4.93 
1  604 

2.15 
1  31 

Lime  

1.075 

0.18 

analyses  of  the  various  metallurgical  products  produced  by 
such  a  procedure.  While  the  several  columns  are  not  all 
strictly  comparable,  they  are  near  enough  so  to  give  a 
good  idea  of  the  character  of  the  products.  Had  not  the 
matte  made  been  charged  back  or  broken  down  before 
skimming  we  should  have  had  another  product  running 
about  26  per  cent  silver  and  0.02  per  cent  gold. 

It  is  interesting  to  note  the  relative  concentrations  of  the 
various  elements  in  the  different  products — the  slimes  pro- 
duce roughly  25  to  30  per  cent  first  slag,  15  per  cent  second, 
slag,  30  per  cent  dore,  and  20  per  cent  flue  dust,  plus  volatile 
losses  and  minus  fluxes  and  furnace  material.  Perhaps  the 
most  striking  thing  is  the  "  spreading "  of  the  silver-gold 
ratio,  nearly  all  the  gold  going  directly  into  the  dore  while 
the  silver  spreads  over  the  various  by-products.  This  ratio 
as  the  figures  stand  is  as  follows:  Slimes,  52.2;  first  slag, 
98.4;  second  slag,  806;  dore,  40.6;  flue  dust,  293. 

The  gold  content  of  the  slags  is  quite  satisfactory  metal- 
lurgically,  but  the  silver  is  entirely  too  high.  When  the 


BY-PRODUCTS  113 

slag  is  reasonably  fusible,  it  is  possible  by  desilverization, 
consisting  of  a  melting  and  collection  of  the  suspended 
values,  to  bring  the  silver  down  to  0.5  per  cent,  but  as 
these  slags  are  all  retreated  in  the  anode  furnace  this  seldom 
pays. 

The  elements  which  do  not  readily  form  fusible  slags 
and  therefore  float  undigested  on  the  surface  of  the  bath  and 
those  (such  as  lead,  iron,  and  antimony)  which  on  the  con- 
trary form  slags  very  readily,  tend  to  concentrate  in  the  first 
slag.  It  is  such  an  impure  and  unpromising  material  that  it 
does  not  offer  much  chance  for  the  recovery  of  by-products. 
Were  the  nickel  out  of  the  way  it  could  be  reduced  to  an 
impure  antimonial  lead. 

The  second,  or  soda  slag  really  represents  impurities  sub- 
sequently blown  out  of  the  base  bullion  and  matte  formed 
when  the  slimes  were  melted.  Here  we  have  a  decided 
concentration  of  tellurium,  an  element  which  clings  tena- 
ciously to  silver.  Since  sodium  salts  are  generally  soluble, 
over  half  the  total  contents  of  this  slag  can  be  leached  out 
with  hot  water.  This  is  the  starting  point  for  the  recovery 
of  tellurium. 

The  small  quantity  of  nitre  slag  made  when  fining  the. 
dore  is  charged  back  into  the  furnace  with  the  next  lot  of 
slimes. 

The  flue  dust  shows  a  remarkable  concentration  of  sele- 
nium and  it  is  from  this  source  that  the  market  demand 
is  supplied.  Elemental  selenium  may  be  readily  prepared 
from  this  dust  in  several  ways — by  fractional  sublimation, 
by  leaching  followed  by  electrolytic  precipitation,  or  by 
leaching  followed  by  chemical  reduction.  The  last-named 
method  is  perhaps  the  most  satisfactory.  The  leaching  is 
generally  conducted  in  the  presence  of  hydrochloric  acid 
and  a  strong  oxidizing  agent  such  as  a  chlorate  and  prefer- 
ably exposed  to  sunlight.  The  resulting  solution  is 
strengthened  in  free  hydrochloric  acid  in  order  to  hold  back 
the  tellurium,  and  the  selenium  is  then  precipitated  as  a 
red  powder  by  sulphur  dioxide  gas.  This  powder  is  melted 
and  skimmed  to  a  bright  face,  when  it  is  cast  into  cakes  as 


114  COPPER  REFINING 

the  black  variety.  If  desired,  the  powder  can  be  dried  in 
an  oven,  the  heat  changing  it  to  the  black  variety  and  the 
baked  product  then  ground  for  the  market.  Selenium  often 
runs  99.5  per  cent  Se  and  the  chief  impurities  are  lead  sul- 
phate, arsenic,  antimony  and  tellurium.  The  lead  content 
is  the  most  difficult  to  control. 

Parting  the  Dore. — We  have  too  well  recognized  methods 
available  by  which  to  part  gold  and  silver — the  sulphuric 
acid  and  the  electrolytic.  Both  are  in  use  and  each  has  its 
advocates  but  the  latter  has  come  into  more  general  use 
chiefly  on  account  of  its  greater  cleanliness. 

In  the  acid  method,  the  dore  is  cast  into  slabs  resembling 
waffles,  the  indentations  giving  additional  surface  for 
corrosion.  The  process  is  very  simple,  being  the  production 
of  silver  sulphate  by  the  decomposition  of  strong,  boiling, 
sulphuric  acid  by  silver,  sulphur  dioxide  being  given  off. 
The  gold  is  left  undissolved  and  is  melted  in  a  crucible, 
while  the  silver  sulphate,  which  is  readily  soluble  in  excess 
of  free  acid,  is  precipitated  upon  copper  sheets  to  form 
sponge  silver  which  is  washed  and  melted.  The  copper 
sulphate  solution  is  sent  to  the  copper-refining  tanks. 

The  plant  required  is  inexpensive — needing  only  a 
couple  of  cast-iron  pots,  a  lead-lined  tank  and  a  crucible 
furnace — and  the  tie-up  of  values  is  short,  being  about  48 
hours.  On  the  other  hand  the  process  is  dirty;  hot  sul- 
phuric acid  is  hard  to  handle;  and  the  fine  silver  cannot  be 
brought  below  a  third  of  an  ounce  in  gold.  Also,  if  the 
dore*  is  high  in  gold,  the  acid  attacks  it  slowly  and  it  is 
necessary  to  inquart  it. 

The  hot  liquor  from  the  kettles  runs  about  600  grams  per 
litre  silver  and  twice  as  much  free  acid.  This  has  to  be 
diluted  with  water  before  the  silver  is  precipitated,  and  the 
tank  house  must  be  able  to  absorb  this  free  acid.  The 
kettles  are  fired  up  and  boiled  until  no  solid  pieces  of  dore 
can  be  felt  with  a  stirring  rod.  This  takes  about  eight 
hours.  The  liquor  is  drawn  off  to  a  heated  settling  kettle  to 
free  it  from  entrained  gold,  fresh  acid  is  added,  and  then 
given  an  additional  boil  to  free  the  gold  of  silver.  By 


BY-PRODUCTS  115 

washing  the  gold  precipitate  thoroughly  before  adding  this 
acid  it  is  possible  to  make  995  gold  sand  directly  in  the 
kettle.  The  silver  sponge  should  run  998  or  better. 

The  electrolytic  process  consists  of  the  electrolysis  of 
dore  anodes  in  a  copper-silver  nitrate  electrolyte.  Three 
methods  have  been  employed — the  Thum-Balbach,  the 
Moebius  and  the  Whitehead.  The  first  method  uses  hori- 
zontal carbon  cathodes  laid  as  the  floor  of  a  shallow  earthen- 
ware tank,  the  anodes  being  carried  in  a  basket  within  a 
light  cotton  bag  to  collect  the  gold  slimes.  At  an  efficiency 
of  90  per  cent,  3.5  volts  per  cell  is  equivalent  to  a  recovery  of 
about  32  oz.  silver  per  kilowatt-hour.  The  current  density 
is  limited  to  about  35  amperes  per  square  foot  by  the  heat- 
ing of  the  electrolyte,  as  a  hot  solution  means  a  high  loss  of 
free  nitric  acid.  The  silver  crystals  have  to  be  frequently 
removed  in  order  to  prevent  treeing,  which  will  reach  and 
puncture  the  gold  bag. 

The  electrolyte  consists  of  about  4  per  cent  copper  as 
nitrate,  2  per  cent  silver  as  nitrate,  and  0.1  per  cent  free 
nitric  acid.  Copper  enters  from  the  dore  and  silver  dis- 
solves in  a  certain  excess,  nitric  acid  being  regularly 
added,  but  the  entrained  liquor  in  the  silver  sponge 
removed  holds  the  balance  steady.  The  wash  waters 
from  this  sponge  are  not  returned  to  the  parting  tanks. 
There  is  no  circulation  of  the  electrolyte  and  each  cell  is 
a  complete  unit. 

The  silver  crystals  are  readily  washed  with  water  and  if 
the  copper  in  the  electrolyte  is  low,  silver  very  close  to  1000 
fine  can  be  produced.  As  this  is  not  desired,  however,  the 
copper  is  allowed  to  build  up  to  a  maximum  of  perhaps  6  per 
cent  in  the  electrolyte  before  any  deliberate  removals  are 
made. 

There  is  no  anode  scrap  as  the  remnants  of  anodes  are 
left  in  place  on  the  horizontal  trays  until  they  are  entirely 
consumed.  Cathode  silver  runs  0.1  oz.  per  ton  in  gold  or 
even  better. 

The  gold  mud  is  periodically  removed  and  boiled  in 
strong  sulphuric  and  in  nitric  acids  to  remove  the  silver. 


116  COPPER  REFINING 

This  mud  is  also  the  raw  material  for  the  recovery  of 
platinum  and  palladium,  as  mentioned  later. 

The  Moebius  process  substitutes  a  vertical  system  of 
electrodes  which  is  very  much  more  compact  and  yields 
more  silver  per  kilowatt-hour,  due  to  lower  tank  resistance. 
The  silver  crystals  drop  to  the  bottom  of  the  tank,  in  some 
plants  being  knocked  from  the  cathode  by  a  simple  system 
of  swinging  wooden  rakes.  The  anodes  are  inclosed  in 
sacks.  Against  the  cheaper  first  cost  and  somewhat  lower 
power  required,  there  is  more  care  required  to  prevent  gold 
leaks,  the  cell  is  not  so  open  to  constant  inspection,  and 
there  is  a  small  amount  of  anode  scrap  made. 

The  third  system,  which  I  have  called  the  Whitehead 
plan,  has  not  come  into  general  use,  although  it  has  been 
more  or  less  extensively  employed  at  some  of  the  govern- 
ment plants.  It  uses  gelatin  to  obtain  a  coherent  deposit  to 
be  stripped  from  silver  cathodes  and  duplicates  in  many 
ways  the  conditions  found  in  a  copper  refining  cell. 

As  compared  with  the  sulphuric  acid  method  of  parting, 
the  products  are  more  pure,  there  is  less  danger  of  metal 
loss,  in  a  large  plant  the  operating  costs  are  somewhat 
lower,  and  but  little  by-product  liquor  is  made.  On  the 
other  hand  the  first  cost  of  plant  is  much  greater,  in  a  small 
operation  the  operating  costs  are  no  lower,  while  the  metal 
tie-up  is  at  least  50  per  cent  more. 

If  platinum  and  palladium  are  present  in  the  dore  we 
have  the  choice  between  electrolytic  and  chemical  recovery. 
If  the  gold  slimes  from  the  parting  are  boiled  to  dissolve 
their  silver  in  sulphuric  acid,  these  metals  remain  with  the 
gold  which  can  then  be  electrolyzed  by  the  Wohlwill  process 
in  a  chloride  electrolyte;  if  nitric  acid  is  used  in  the  gold 
boiling,  the  platinum  and  palladium  go  chiefly  into  solu- 
tion and  after  a  somewhat  tedious  treatment  may  be 
recovered  as  double  chlorides  with  ammonium.  The  plati- 
num is  thrown  out  singly  by  sal  ammoniac  from  a  reduced 
solution;  after  oxidation  the  palladium  may  be  similarly 
precipitated.  Metallic  sponges  are  obtained  upon  ignition 
of  these  precipitates.  As  these  sponges  are  difficult  to 


BY-PRODUCTS  117 

melt  the  metal  is  usually  marketed  as  a  crude  powder. 
The  separation  between  the  two  metals  is  quite  sharp  and 
a  product  98  per  cent  or  better  is  readily  obtained. 

The  Wohlwill  process  has  two  great  objections:  it  ties  up 
a  very  valuable  metal,  and  it  exposes  it  to  possible  theft. 
In  ordinary  practice  the  gold  does  not  appear  in  solid  yel- 
low metal  until  it  is  almost  ready  to  ship.  Crude,  dirty 
products  are  hard  to  sell  and  do  not  present  the  same  temp- 
tation to  laborers  handling  them.  On  the  other  hand  the 
electrolytic  process  is  cleaner,  as  in  the  case  of  silver  part- 
ing, with  the  result  that  most  government  plants  use  elec- 
trolytic and  most  private  plants  chemical  extraction.  As 
before  stated,  the  government  does  not  pay  for  platinum 
contents  of  deposits;  one  reason  for  this  is  that  the  assay 
method  for  accurate  results  is  very  tedious  and  expensive 
and  its  application  would  result  in  great  delays  in  making 
settlements. 

Slimes  Process  Development. — The  progress  already 
made  with  the  metallurgical  problems  presented  by  copper 
anode  slimes  may  be  seen  from  Figs.  25  and  26  which  are 
flow  sheets  representing  practice  twenty  years  ago  and  to- 
day. We  may  now  consider  the  various  points  at  which 
the  second  process  may  be  attacked. 

Screens. — The  fact  that  a  certain  part  of  the  copper  car- 
ried in  the  slimes  may  be  removed  by  simple  screening  sug- 
gests the  possibility  of  further  separation  by  mechanical 
means. 

The  first  point  of  attack  here  is  against  the  size  of  the 
particle — could  the  slimes  be  classified  hi  any  way  by 
screening?  The  practice  at  one  plant  is  to  screen  out  no- 
dules and  large  pieces  by  passing  the  slimes  through  a  4- 
mesh  screen,  and  then  through  a  70-mesh  screen,  which 
takes  out  float  slime  such  as  was  discussed  in  Table  24 
in  the  last  chapter.  Changing  the  size  of  this  fine  screen 
and  then  analyzing  the  slimes  gave  the  results  shown  in 
Table  38,  below,  which  indicate  the  hopelessness  of  such  a 
classification. 


118 


COPPER  REFINING 


Anode  Furnaces 


, 

Tank  House 


Lead 


Slimes 

1st 

Irtn 

1st  Cupel 

Skim 

2nd  Cupel 

I  Liguor 

J- [Copper   . 

Iron  Tank  , — —>•    Anode 
Furnaces 


Lead 


Tank  House 


Tank  House 


Cupel 


T  Gold 
Market 


FIG.  25.— Slimes  process— 1900. 


TABLE  38. — ANALYSES  OF  SLIME  THROUGH  DIFFERENT  SIZE  SCREENS 


Meshes  per  inch 

Per  cent 
copper 

Per  cent 
arsenic 

Per  cent 
antimony 

-     4 

21.3 

6.00 

2.44 

-  80 

19.6 

5.39 

3.10 

-120 

19.4 

5.50 

3.92 

-200 

21.3 

4.30 

3.60 

BY-PRODUCTS 

Slit 


119 


H2S04 


Dore  Furnace 


S"0|  -  *         Wash  Tank 


Crucible 
[Gold 


IGold  I  Platinum    I 

Sand  |  Palladium^ 


Market 


Market 


Market 


Market 


FIG.  26. — Slimes  process— 1915. 


A  screen  analysis  of  slimes  is  given  in  Table  39.  Here 
we  happen  to  have  a  sample  which  has  some  small  cathode 
debris  present  and  the  effect  of  screening  this  out  is  at  once 
apparent  upon  the  copper  assays  and  the  use  of  one  fine 
screen  is  justified.  In  Table  38  the  assay  is  for  the  total 
slimes  passed  by  the  screen  while  Table  39  gives  results  for 
a  true  screen  analysis.  A  70-mesh  screen  will  ordinarily  re- 


120 


COPPER  REFINING 


tain  8  or  9  per  cent  of  the  weight  of  slimes  sent  it  from  the 
4-mesh,  and  this  quantity  may  be  advantageously  returned 
to  the  anode  furnace. 

TABLE  39. — SCREEN  ANALYSIS  OF  SLIMES 


Mesh 
passed 

Mesh 
retained 

Grains 

Per  cent 
weight 

Per  cent 
copper 

Grams 
copper 

Per  cent 
total 
copper 

4 

20 

319.1 

55.6 

40.8 

130.2 

74.3 

20 

40 

101.4 

17.6 

27.0 

27.4 

15.6 

40 

60 

54.0 

9.4 

22.0 

11.9 

6.8 

60 

80 

30.8 

5.4 

13.7 

4.2 

2.4 

80 

100 

14.4 

2.5 

10.6 

1.5 

0.9 

100 

54.3 

9.5 

Not  run 

Total  

574.0 

100.0 

30.6 

175.2 

100.0 

As  the  silver  and  gold  are  presumably  present  in  metallic 
form  and  as  they  have  a  higher  specific  gravity  than  most 
of  the  other  elements  present,  some  gravity  separation 
might  classify  the  product.  Careful  trials  with  a  Wilfley 
table  did  not  effect  sufficient  separation,  however,  to  en- 
able any  process  to  be  built  on  such  a  basis.  It  is  just 
possible  that  the  application  of  recent  flotation  experience 
might  show  something  of  value.  It  must  be  remembered 
that  what  is  necessary  is  not  so  much  a  perfect  separation 
— any  process  will  have  to  work  up  its  intermediate  prod- 
ucts for  precious  metal  values — as  a  scheme  whereby  the 
great  bulk  of  the  silver  and  gold  will  be  put  on  the  market 
promptly,  leaving  the  more  tedious  working  of  other  by- 
products until  the  last.  One  of  the  objections  to  present 
practice  is  that  the  other  elements  are  gradually  removed 
until  the  pure  silver  and  gold  are  obtained,  thereby  holding 
the  entire  value  until  the  very  end  of  a  long  process. 

Oxidizing  the  Copper. — The  next  step  in  the  present 
process  is  oxidation  of  the  copper.  It  has  already  been 
pointed  out  that  the  presence  of  any  quantity  of  copper 
in  the  dore-furnace  charge  makes  a  serious  amount  of  by- 


BY-PRODUCTS  121 

product  matte  high  in  silver.  If  the  copper  is  allowed  to 
stand  at  15  or  20  per  cent  and  the  slimes  charged  directly 
into  the  furnace,  instead  of  obtaining  30  per  cent  first  slag, 
15  per  cent  second  slag  and  30  per  cent  dore*  we  should  get 
perhaps  35  per  cent  slag,  25  per  cent  matte  and  only  20 
per  cent  dore,  a  third  of  the  silver  being  tied  up  in  the  matte. 
The  desirability  of  the  awkward  oxidation  step  is  thus 
fully  demonstrated. 

In  the  endeavor  to  simplify  this  operation,  considerable 
work  has  been  done  experimentally  with  ferric  sulphate  as 
an  oxygen  carrier  using  the  well  known  reactions : 

Cu  +  Fe2(SO4)3  =  CuS04  +  2FeSO4 
and 

2FeSO4  +  CuSO4  +  current  =  Fe2(SO4)3  +  Cu. 

A  diaphragm  cell  is  all  that  is  necessary  to  carry  out 
this  process  and  chemically  it  is  very  pretty.  The  fact 
that  it  has  not  been  successfully  applied  even  in  the 
hands  of  those  skilled  in  electrolysis  bears  witness  to 
the  great  difficulties  interposed  by  a  diaphragm  cell  in 
an  acid  process.  Later  work  such  as  that  of  Hybinette 
in  nickel  or  the  use  of  a  chemical  retarder1  may  reopen  this 
field. 

Leaching. — The  next  step  of  dissolving  the  oxidized 
copper  from  the  slimes  is  true  leaching,  and  leads  to  the 
consideration  of  the  possibility  of  dissolving  the  bulk  of 
the  slimes  at  once  instead  of  making  this  a  mere  preparatory 
step  to  furnacing. 

The  first  suggestive  point  to  be  noted  is  that  in  any  of  the 
oxidation  processes  it  is  easy  to  carry  the  action  to  a  point 
where  large  quantities  of  silver  may  be  taken  into  dilute 
sulphuric  acid  solution  as  sulphate.  It  is  further  known 
that  by  careful  temperature  regulation,  a  very  high  grade 
selenium  product  may  be  obtained  in  the  flue.  On  the 
other  hand,  anything  but  an  incipient  roast  tends  to  frit 
the  easily  melted  slimes,  spoiling  the  physical  condition 

1  Addicks,  U.  S.  patent  No.  1138921. 


122  COPPER  REFINING 

for  subsequent  leaching  and  forming  undesirable  chemical 
compounds  of  copper,  arsenic,  etc.  Roasting  therefore 
offers  possibilities  which  are  limited  by  difficulties  in 
practical  operation. 

A*  second  method  of  attack  is  to  digest  the  raw  slimes  in 
boiling  strong  sulphuric  acid.  This  readily  dissolves 
nearly  the  entire  content  and  yields  a  very  foul  solution  of 
sulphates  and  an  impure  gold  residue.  Fortunately  silver 
forms  one  of  the  very  few  insoluble  chlorides  and  the  plan 
would  be  to  precipitate  the  silver  with  salt  and  reduce  the 
chloride  to  metallic  silver. 

We  have  now  a  plan  which  would  promptly  yield  the 
bulk  of  both  the  silver  and  the  gold  for  marketing  while 
leaving  the  less  valuable  constituents  behind  for  working 
up  into  various  other  by-products.  Such  a  plan  may  yet 
develop  commercially  but  there  are  again  some  serious 
practical  problems  to  be  overcome.  In  the  first  place  a 
certain  amount  of  flour  gold  is  formed  and  it  is  very  difficult 
to  keep  this  away  from  the  silver  chloride  precipitate.  If 
the  latter  is  not  free  from  gold  it  has  to  be  parted,  incurring 
corresponding  delay  and  expense.  Silver  sulphate  is  not  a 
readily  soluble  salt  and  a  considerable  bulk  of  valuable 
liquor  has  to  be  handled.  The  reduction  of  silver  chloride 
can  be  done  electrolytically  or  chemically  but  even  that  is 
not  complete  and  the  question  of  small  quantities  of  chloride 
being  volatilized  during  melting  has  to  be  considered. 

On  the  gold  side  we  have  a  sludge  carrying  lead  and 
antimony  as  well  as  the  gold,  and  this  involves  fluxes  and 
slags  in  the  refining. 

The  other  elements  are  collected  in  a  bulky,  dilute  solu- 
tion carrying  sulphates  and  chlorides  and  any  simple 
separation  is  a  chemical  problem  of  no  mean  order. 

Smelting. — Another  possibility  lies  in  directly  melting 
the  raw  slimes  in  a  furnace  with  separation  of  slag  and  matte 
leaving  most  of  the  gold  and  half  to  two-thirds  of  the  silver 
in  a  dore  bullion  which  can  be  quickly  fined  and  parted. 
The  slag  can  be  returned  to  the  anode  furnace  as  at  present 
and  the  matte  subjected  to  some  special  process. 


BY-PRODUCTS  123 

The  direct  blowing  of  this  matte'  in  a  small  Bessemer 
converter  is  the  most  obvious  method.  We  would  get  a 
selenious  oxide  fume  and  a  copper-silver  blister  as  the 
products,  which  would  be  quite  satisfactory  metallurgically. 

The  difficulty  lies  in  the  lack  of  courage  to  face  the 
possible  silver  losses.  Now  that  the  Cottrell  process  has 
been  so  successfully  applied  to  silver  flue  losses1  this  might 
be  reconsidered. 

Another  method  is  to  pulverize  this  matte  and  leach  the 
values,  pre-roasting  if  desired.  This  really  means  leaching 
out  the  copper  and  we  have  simply  rearranged  the  steps  in 
the  original  process  so  as  to  more  promptly  release  part  of 
the  values. 

Smelting  has  the  advantage  that  a  fab*  separation  of  other 
impurities  is  made  in  concentrated  form,  easily  stored  for 
retreatment. 

Parting. — This  step  is  in  a  very  satisfactory  state  of 
development  and  but  little  improvement  in  methods  is  to 
be  looked  for. 

1  Aldrich,  Transactions  American  Electrochemical  Society,  Vol.  xxviii,  p. 
119. 


CHAPTER  VIII 
FURNACE  REFINING 

In  the  early  days  copper  was  produced  by  a  succession 
of  roasting  and  fusing  operations  based  upon  the  fact 
that  copper  sulphide  when  melted  with  copper  oxides 
reacts  to  form  sulphur  dioxide  gas  and  metallic  copper. 
This  was  the  basis  of  the  celebrated  Welsh  process.  Then 
it  was  found  that  if  the  operation  was  so  conducted  that  a 
small  metal  fall  was  obtained  first,  these  " bottoms" 
would  contain  most  of  the  impurities,  including  gold  (silver 
was  frequently  separated  by  a  sulphatizing  roast  and  sub- 
sequent wet  extraction),  and  the  balance  of  the  copper 
produced  would  be  correspondingly  purer.  This  is  the 
characteristic  "best  selected"  process  used  in  England  and 
the  basis  of  the  "Argo"  process  for  separating  gold  from 
copper. 

While  converting  has  supplanted  roasting  and  elec- 
trolysis reverberatory  refining  for  the  great  bulk  of  the 
world's  copper,  although  the  melting  of  Lake  mineral  in 
Michigan  and  various  small-scale  operations  abroad  are 
notable  exceptions,  the  electrolytic  process  has  retained 
two  reverberatory  steps,  the  anode  and  refining  furnace 
operations,  in  the  metallurgical  scheme. 

Blister  copper  produced  directly  by  the  converter  does 
not  make  desirable  anodes,  although  special  conditions 
have  at  times  made  their  use  profitable.1  For  the  general 
case  the  principle  has  been  well  established  that  work 
which  can  be  done  in  the  anode  furnace  should  not  be  done 
in  the  tank-house.  The  guiding  principle  in  tank-house 
operation  is  to  do  whatever  is  necessary  to  obtain  uni- 
formity of  operating  conditions,  and  converter  anodes 
give  trouble  unless  a  very  pure  blister  is  being  treated. 

1  Burns,  Trans.  A.  I.  M.  E.,  vol.  xlvi,  p.  711. 

124 


FURNACE  REFINING  125 

The  function  of  the  anode  furnace  is  to  make  a  perfect 
anode  casting  in  which  the  sulphur  has  been  thoroughly 
removed  and  other  oxidizable  impurities  slagged  off  as  far 
as  possible. 

The  refining  furnace  serves  merely  to  put  the  cathodes 
into  desired  physical  shape  and  beyond  the  elimination  of 
traces  of  sulphur  and  adjustment  of  " pitch"  has  ordinarily 
no  chemical  function.  Nevertheless  the  methods  of  opera- 
tion of  the  two  furnaces  are  identical  in  many  respects 
and  the  discussion  of  the  general  principles  involved  is 
applicable  equally  to  both. 

Just  as  many  of  the  operations  in  smelting  are  based 
upon  the  marked  affinity  between  copper  and  sulphur, 
so  the  basis  of  furnace  refining  is  the  relative  weakness  of 
the  affinity  between  copper  and  oxygen.  The  general 
plan  is  therefore  to  oxidize  the  impurities  the  oxides  of 
which  will  then  either  escape  by  sublimation  or  float  on 
the  surface  of  the  bath,  from  which  they  may  be  skimmed 
or  slagged.  According  to  this  method  it  should  be  possible 
to  remove  all  elements  less  noble  than  copper,  leaving 
simply  silver,  gold  and  the  platinum  group  alloyed  with 
the  copper.  The  same  principle  of  resistance  to  oxidation 
is  reversed  in  the  tank  house,  the  copper  oxidizing  to  sul- 
phate at  the  anode  and  the  more  noble  metals  remaining 
behind  in  the  slimes.  Theoretically  we  have,  therefore,  a 
perfect  separation  of  the  copper,  the  impurities  and  the 
values. 

In  practice  the  system  does  not  work  out  so  perfectly, 
as  mass  action  comes  into  play.  It  is  quite  easy  to  remove 
the  bulk  of  most  oxidizable  impurities  in  the  anode  furnace, 
but  as  the  amount  present  becomes  less  and  less  it  is  in- 
creasingly difficult  to  make  further  removals,  and  carried 
to  absurdity  the  entire  charge  would  be  slagged  in  order  to 
remove  the  last  traces.  Then  some  elements  such  as 
arsenic  readily  form  direct  compounds  with  the  more 
basic  metals,  and  these  cannot  readily  be  broken  up.  The 
degree  of  elimination  of  impurities  depends  therefore  upon : 
(1)  the  amount  of  impurity  present;  (2)  its  affinity  for 


126  COPPER  REFINING 

oxygen;  (3)  its  affinity  for  copper,  and  (4)  the  extent  to 
which  scorification  is  pushed. 

Scorification  consists  of  using  an  element  as  an  oxygen 
carrier.  In  this  case  the  principal  scorifying  agent  is 
copper  in  the  form  of  cuprous  oxide.  The  process  is  to 
oxidize  with  an  air  blast  some  of  the  impurities  directly 
and  a  part  of  the  copper  by  air.  The  cuprous  oxide  formed 
dissolves  in  the  bath  and  penetrates  to  all  parts  of  it, 
reacting  with  elements  which  have  a  greater  affinity  than 
copper  for  oxygen  and  reducing  out  an  equivalent  amount 
of  copper.  These  basic  oxides  further  react  with  any  acid 
material,  such  as  silica  in  the  furnace  walls,  that  may  be 
present  and  the  resulting  slag  is  skimmed  off,  leaving  a 
purified  bath  holding  a  considerable  quantity  of  cuprous 
oxide  in  solution. 

The  reducing  stage  necessary  to  throw  back  the  c'uprous 
oxide  is  now  accomplished  by  generating  a  hydrocarbon 
gas  by  the  destructive  distillation  of  hard  wood  poles 
thrust  beneath  the  surface  of  the  bath,  which  is  covered 
by  a  blanket  of  carbonaceous  material  to  protect  it  from 
any  further  oxidation.  The  various  steps  will  now  be 
taken  up  in  detail  in  the  order  in  which  they  occur. 

Charging. — A  reverberatory  furnace  is  used  for  this 
work  operated  on  a  twenty-four-hour  cycle.  Starting 
with  an  empty  furnace  the  first  operation  is  charging.  In 
the  early  days  this  was  accomplished  by  hand,  one  pig  or 
cathode  being  charged  at  a  time.  Two  men  would  lift  a 
pig  to  the  end  of  a  peel  held  in  position  at  a  charging  door  by 
a  third  man.  This  peel  was  a  long  heavy  steel  bar  flattened 
like  a  paddle  at  one  end.  The  three  men  would  then  thrust 
the  peel  into  the  furnace  and  flip  the  pig  off  at  the  desired 
location. 

Various  mechanical  devices  were  tried  to  improve  this 
practice,  but  for  a  long  time  the  only  successful  innovation 
was  the  use  of  air  lifts  and  tongs  in  placing  the  pigs  on  the 
peel.  As  furnaces  gradually  increased  in  size  the  time 
required  to  charge  them  became  inordinate,  and  not  only 
was  the  twenty-four-hour  cycle  badly  disarranged,  but  the 


FURNACE  REFINING  127 

furnace  cooled  off  badly  during  the  delay.  These  diffi- 
culties led  to  trials  of  charging  cranes  developed  in  the 
open-hearth  steel  industry,  but  they  were  at  first  unsuccess- 
ful, due  to  a  lack  of  appreciation  of  fundamental  differences 
between  the  two  problems.  The  open-hearth  furnace  has  a 
small  charge  simply  thrown  in  it  while  the  copper  reverbera- 
tory  by  the  old  practice  was  piled  with  pigs  clear  to  the  roof. 
The  first  result  of  mechanical  charging  was  an  enormous 
loss  in  cold  capacity  which  more  than  offset  the  other 
advantages  gained.  This  failure  checked  further  develop- 
ment along  these  lines  for  some  time  until  a  special  form  of 
peel  consisting  of  a  fork  with  a  special  pushing  device1 
was  devised  to  replace  the  upsetting  shovel  of  the  old  type. 
This  crane  met  with  immediate  success,  and  it  was  found 
possible  to  charge  5000  Ib.  at  a  time  at  a  rate  of  at  least 
200,000  Ib.  an  hour  and  lose  nothing  in  cold  capacity  as 
compared  with  the  hand  method,  while  at  the  same  time 
charging  a  200-ton  furnace  as  quickly  as  the  old  50  and 
75-ton  units  had  formerly  been  handled. 

The  main  thing  in  charging  is  to  have  material  which  will 
stack  well.  This  means  that  pigs  should  have  a  good  blister 
finish  and  not  be  overblown  and  that  cathodes  should  be 
reasonably  smooth.  When  this  is  not  the  case  individual 
charges  after  piling  in  the  furnace  will  spill  and  take  up 
much  more  space. 

The  next  development  along  this  line  was  the  use  of  the 
crane  for  recharging  after  melting  the  first  batch.  In  this 
way  a  certain  molten  tonnage  in  excess  of  the  normal  cold 
capacity  of  the  furnace  can  be  obtained. 

Then  it  was  found  that  anode  scrap  could  be  recharged 
to  a  certain  extent  while  the  metal  was  being  poured. 
Finally  cathodes  were  successfully  charged  while  the  cop- 
per was  being  poured.2  This  will  be  referred  to  again 
later  on. 

In  this  way  the  limit  to  size  has  been  practically  elimi- 
nated as  mechanical  charging  will  rapidly  take  care  of  any 

1  Prosser  and  Ladd,  U.  S.  Patent. 

2  Addicks  and  Marks,  U.  S.  Patent  No.  980584. 


128  COPPER  REFINING 

size  of  furnace  which  can  be  built  and  the  cost  has  been 
brought  down  greatly  as  the  hand  method  exposed  the 
crew  to  very  severe  heat. 

Melting. — The  next -step  in  refining  is  the  melting  of  the 
charge.  This  consists  merely  of  burning  as  much  fuel  as 
the  furnace  can  take  care  of.  Copper  is  a  good  conductor 
of  heat  and  the  piled-up  charge  through  which  the  hot 
gases  have  to  filter  absorb  heat  very  rapidly.  The  melting 
is  divided  into  three  stages:  (1)  " softening,"  when  the  hot 
mass  collapses;  (2)  " coming  flat,"  when  the  molten  metal 
covers  the  softened  mass;  and  (3)  " coming  afloat,"  when 
the  balance  remaining  on  the  bottom  is  absorbed. 

It  has  been  found  that  a  semi-bituminous  coal  running 
about  20  per  cent  in  volatile  matter  gives  a  desirable  length 
of  flame  for  this  work.  It  is  most  essential  that  the  coal 
should  not  form  a  fusible  ash  as  deep  fires  are  carried,  and 
unless  they  are  free  burning  a  great  deal  of  delay  is  en- 
countered. In  a  refining  furnace  the  old  practice  has  been 
to  fire  heavily  during  the  early  stages  of  the  refining  and 
carry  a  bed  of  coke  for  heat  during  the  later  part  so  as  to 
avoid  sulphur  from  the  fuel  reaching  the  copper.  With  the 
present  large  charges,  and  particularly  when  recharging, 
while  pouring  it  has  been  found  necessary  to  fire  more 
steadily  and  without  any  apparent  untoward  effect. 

Flapping. — When  the  charge  is  afloat  a  button  is  taken  in 
a  "say  ladle,"  and  if  this  throws  a  "worm,"  indicative  of 
sulphur,  a  pole  is  introduced  to  agitate  the  bath,  which  is 
exposed  to  the  air  until  a  button  sample  shows  a  surface 
which  is  unbroken  on  cooling.  This  practice  has  a  bearing 
on  "overpoled"  charges  of  refined  copper  to  be  taken  up 
later. 

The  system  of  using  buttons  for  samples  is  very  old,  but 
still  stands  as  the  best  method  of  controlling  furnace  re- 
fining. A  small  ladle  with  a  cup  about  an  inch  and  a  half 
in  diameter  is  warmed  by  holding  over  the  bath  for  a 
moment,  then  used  as  a  rabble  to  sweep  a  clean  face  at  the 
spot  where  the  sample  is  to  be  taken  and  then  plunged 
beneath  the  surface  and  quickly  withdrawn.  The  sample 


FURNACE  REFINING  129 

thus  obtained  is  allowed  to  "set,"  when  it  may  be  dumped 
into  water  to  cool  for  handling.  Both  the  surface  and  the 
fracture  upon  nicking  with  a  chisel  and  breaking  in  a  vise 
give  clear  indications  as  to  the  state  of  the  bath  of  copper. 
When  sulphur  is  present  a  small  volcano  forms  on  cooling, 
which  throws  out  a  part  of  the  metal  which  twists  around  as 
it  chills  much  like  a  worm.  When  excess  oxygen  is  present 
the  surface  shows  a  deep  depression,  and  when  the  bath  is 
nearly  saturated  with  cuprous  oxide  this  depression  will 
break  inward  into  a  cavity.  Copper  at  desired  "pitch" 
for  pouring  will  show  a  full  rounding  slightly  wrinkled 
surface.  Overpoled  copper  begins  to  "spew"  and  show 
the  above-mentioned  sulphur  indications. 

In  the  same  way  the  condition  of  affairs  may  be  noted  by 
the  color  and  structure  of  the  fracture.  When  copper  is 
saturated  with  cuprous  oxide  (about  6  per  cent  Cu2O, 
beyond  which  point  further  oxidation  will  only  cause  useless 
slagging)  the  fracture  shows  a  marked  cubical  structure, 
brick-red  in  color.  As  the  oxygen  is  poled  out  the  struc- 
ture becomes  more  fibrous  and  the  color  brighter  until  we 
reach  the  lustrous  silky  texture  corresponding  to  properly 
refined  copper. 

These  simple  tests  enable  a  practiced  eye  to  follow  the 
refining  operations  with  very  satisfactory  accuracy,  and 
consequently  metallographic  and  other  methods  have  not 
come  into  use  to  any  extent. 

When  the  sulphur  has  been  poled  out  the  flapping  or 
blowing  begins.  Originally  a  rabble  was  swung  from  a 
chain  at  the  furnace  door  and  the  surface  of  the  metal 
flapped  or  broken  into  spray  by  swinging  this  back  and 
forth  with  free  admission  of  air  at  all  doors.  This  is  hot 
and  fatiguing  work,  and  as  the  size  of  furnace  charges  in- 
creased it  again  became  necessary  to  find  some  mechanical 
substitute.  This  was  accomplished  by  blowing  compressed 
air  beneath  the  surface  of  the  bath,  sending  up  a  fountain  of 
molten  metal  instead  of  a  mere  splash  and  at  the  same  time 
thoroughly  stirring  the  bath  itself.  The  great  objection  to 
this  procedure  has  from  the  beginning  been  the  consequent 


130  COPPER  REFINING 

damage  to  the  furnace  structure.  The  fountain  is  so  violent 
that  more  or  less  of  the  spray  reaches  the  roof  of  the  fur- 
nace and  the  metal  highly  charged  with  cuprous  oxide  avidly 
attacks  the  silica  brick  wherever  it  touches  it.  The  large 
modern  furnaces  tend  toward  excessive  repair  costs  as  one 
of  their  disadvantages,  and  this  corrosive  action  aggravates 
a  situation  already  bad.  When  basic  or  neutral  material 
such  as  chrome  brick  is  substituted  for  acid  brick  construc- 
tion, this  trouble  disappears;  otherwise  constant  attention 
must  be  given  to  restraining  the  men  in  control  of  the  air 
blast  who  naturally  tend  to  over-vigorous  application  in 
order  to  gain  time.  Another  remedy  would  be  the  use  of 
low  pressure  air,  which  would  not  give  such  violent  agita- 
tion, but  the  standard  80  Ib.  per  square  inch  general  service 
air  system  around  the  plant  is  generally  utilized. 

Chemically  the  method  is  very  efficacious  and  the  largest 
bath  can  be  thoroughly  saturated  with  cuprous  oxide  in  a 
couple  of  hours  at  small  expense.  The  iron  pipe  used  gradu- 
ally dissolves,  but  the  amount  of  iron  thus  absorbed  by  the 
copper  is  proportionately  so  small  as  to  be  of  no  consequence 
even  in  a  cathode  melting  furnace. 

Skimming. — The  next  operation  is  the  removal  of  the 
slag,  which  is  simply  pulled  off  with  a  rabble  through  the 
skimming  door  into  suitable  pots.  The  skulls  are  later 
broken  up  and  if  the  slag  made  was  reasonably  fluid,  the 
large  amount  of  metallics  always  present  can  be  easily 
sorted  out  and  returned  to  the  furnace.  Often  slag  rich  in 
metallics  is  returned  to  the  next  charge. 

The  general  question  of  slag  formation  requires  con- 
sideration at  length  and  will  be  taken  up  later  on. 

Coking. — When  the  bath  has  been  skimmed  to  a  clean 
face,  charcoal,  coke  or  anthracite  coal  is  spread  over  the 
surface  to  protect  it  from  oxidizing  influences  during  the 
subsequent  poling  operation.  At  first  charcoal  was  ex- 
clusively used,  particularly  when  melting  cathodes,  on  ac- 
count of  its  low  ash  and  freedom  from  sulphur.  Then  the 
much  cheaper  low-sulphur  coke  crushed  to  suitable  size 
was  substituted,  reserving  charcoal  merely  for  the  cover- 


FURNACE  REFINING  131 

ing  on  the  pouring  ladle  between  the  furnace  and  the 
casting  machine,  where  the  finishing  touches  are  always 
put  in  adjusting  the  oxygen  content  of  the  product.  Fi- 
nally buckwheat  coal  has  been  largely  used  on  anode 
charges  as  the  cheapest  form  of  carbon  for  such  work; 

As  this  blanket  of  floating  fuel  burns  away  it  has  to  be 
renewed  so  that  no  bare  surface  be  exposed  to  possible 
oxidation. 

Poling. — As  regards  poling,  as  the  reduction  of  the  excess 
cuprous  oxide  is  called,  little  or  no  change  has  been  made 
from  early  practice.  As  the  furnaces  have  increased  in  size 
larger  and  longer  poles  have  been  used,  but  the  method  of 
forcing  one  end  of  a  pole  below  the  surface  of  the  bath  by 
using  a  chain  tackle  making  an  inverted  lever  against  an 
iron  cross  bar  is  still  in  use.  As  the  pole  is  consumed  it  is 
gradually  fed  in  and  a  new  grip  taken  with  the  chain  until 
the  stick  becomes  too  short,  when  it  is  thrown  into  the 
furnace  as  a  floating  brand.  Green  hard  wood  gives  the 
best  results. 

The  poling  operation  is  a  very  clumsy  one*  The  poles 
are  hard  to  handle  and  bulky  to  store.  When  a  single 
charge  takes  twenty-five  young  trees  to  pole  it  up  it  may 
readily  be  seen  that  the  surrounding  country  will  soon  be 
stripped  of  suitable  timber  and  that  transportation  charges 
make  poling  cost  an  ever-increasing  burden. 

Oil  poling  has  been  tried  with  but  partially  satisfactory 
results.  The  reduction  is  very  rapid  when  fuel  oil  is  blown 
beneath  the  surface  of  molten  copper,  a  few  minutes'  time 
corresponding  to  two  hours'  ordinary  poling.  In  fact,  its 
very  rapidity  of  action  is  one  of  the  objections  to  its  use, 
as  it  is  difficult  to  arrest  the  reducing  action  before  the 
copper  is  "overpoled."  Another  serious  difficulty  attend- 
ant on  the  use  of  oil  wlien  poling  cathode  charges  is  that 
the  copper  absorbs  sulphur  or  other  impurities  from  the 
oil,  which  give  a  very  curious  appearance  to  its  surface 
on  cooling,  and  would  doubtless  make  its  sale  difficult  even 
if  no  objectionable  property  were  imparted  to  it. 

Finally  oil  poling  is  very  variable  in  its  expense.     Appa- 


132  COPPER  REFINING 

rently  it  is  very  easy  to  use  twice  the  amount  to  pole  one 
charge  that  is  necessary  for  another,  the  excess  oil  being 
burned  above  the  surface  of  the  charge. 

These  difficulties  have  prevented  the  general  use  of  direct 
oil  poling.  Nevertheless  it  seems  probable  that  the  time 
will  come  when  gas  poling  will  be  used.  It  is  evident  that 
an  outside  source  of  heat  can  be  used  for  the  destructive 
distillation  of  wood  or  that  oil  or  coal  gas  could  be  made 
suitable  for  the  purpose.  In  the  case  of  wood  the  cheapest 
available  form  of  waste  could  be  utilized  instead  of  valu- 
able young  trees.  Such  a  gas  could  be  readily  controlled  in 
its  action  and  a  gas  made  from  wood  waste  would  require 
no  purification. 

As  the  poling  operation  proceeds  sample  buttons  are 
frequently  taken  for  examination,  and  when  the  desired 
amount  of  reduction  as  shown  by  these  tests  is  nearly  at- 
tained an  ingot  is  poured  and  the  nature  of  the  "set"  sur- 
face on  the  top  of  this  casting  inspected.  When  a  full 
rounding  surface  is  obtained  the  charge  is  ready  to  pour. 
Castings  of  different  sizes  require  a  different  amount  of 
unreduced  oxide  in  order  to  obtain  a  level  surface  in  an  open 
mold.  The  final  adjustment  is  therefore  made  in  the  casting 
ladle,  where  charcoal  is  added  or  allowed  to  burn  thin 
according  to  the  appearance  of  the  set  on  the  castings  being 
made. 

Should  poling  inadvertently  be  allowed  to  continue  too 
long  the  charge  becomes  "overpoled,"  a  condition  the  symp- 
toms of  which  are  much  better  understood  than  the  under- 
lying causes.  The  surface  of  a  casting  made  from  such 
copper  swells  as  it  cools  until  it  breaks  or  "  spews,"  giving 
a  "  worm"  similar  to  that  noted  in  connection  with  sulphur- 
bearing  anode  charges.  While  such  a  condition  can  be 
apparently  rectified  by  a  slight  oxidation  of  the  charge, 
copper  in  this  state  becomes  unmanageable  as  regards  its 
set,  and  such  rapid  vacillations  are  encountered  while  pour- 
ing that  accepted  practice  is  to  flap  the  charge  down  again 
to  "dry"  copper  and  pole  back  to  the  proper  oxygen  con- 
tent, or,  in  short,  to  re-refine  the  entire  charge.  The 


FURNACE  REFINING  133 

unstable  condition  can  be  controlled  by  adding  certain  sub- 
stances, notably  lead,  in  small  quantities,  but  as  mere  traces 
of  lead  make  copper  brittle  the  cure  is  worse  than  the  dis- 
ease. It  is  probable  that  sulphur  plays  some  part  in  this 
peculiar  behavior. 1 

Pouring. — Furnaces  up  to  20  tons  in  capacity  used  to  be 
ladled  by  hand,  a  group  of  five  or  six  men  each  with  a  ladle 
holding  some  15  or  20  Ib.  of  copper  forming  an  endless  chain 
between  the  " ladle  hole"  and  the  molds.  The  latter  were 
made  of  cast  iron  or  copper  placed  over  a  shallow  bosh  and 
the  castings  were  dumped  by  hand  into  the  water  from 
which  ,they  were  fished  with  suitable  tongs.  The  next 
development  was  the  "bull  ladle,"  which  is  still  in  use  at 
some  places.  This  was  simply  a  large  ladle  slung  from  an 
overhead  crawl,  enabling  one  man  to  dip  out  as  much  at  a 
time  as  the  five  or  six  had  done  before. 

Then  came  the  casting  machines,  which  have  developed 
along  several  lines.  Three  general  types  are  in  use — the 
endless  chain  conveyor  first  developed  by  McCoy,  the  well- 
known  Walker  wheel  and  the  variation  of  this  known  as 
the  Clark  machine.  These  are  all  covered  by  various 
patents,  most  of  which,  however,  have  now  expired. 

The  first,  which  has  now  gone  out  of  use  for  wire  bars, 
but  is  still  occasionally  employed  for  anodes  and  more 
recently  in  casting  blister  copper  into  pig,  carried  the  molds 
on  a  link  belt  mechanism  in  front  of  the  spout,  then  lowered 
them  on  inclined  rails  into  a  bosh  of  water  and  then 
dumped  the  castings  as  the  molds  returned  upside  down  on 
the  return  trip  underneath  the  bosh.  The  Walker  wheel, 
as  its  name  implies,  has  a  central  hub  carrying  the  rotating 
mechanism  and  radial  arms  supporting  the  molds.  The 
Clark  machine  differs  from  the  Walker  chiefly  in  that  the 
molds  are  carried  parallel  to  the  radial  arms  while  on  the 
latter  they  are  placed  circumferentially,  and  still  further 
modifications  have  been  made  at  the  new  Great  Falls  plant. 

The  Walker  scheme  has  two  advantages  over  each  of  the 
other  types  due  to  the  way  the  molds  are  placed  when 

1  Skowronski,  Trans,  A.  I.  M.  E.,  vol.  Ix,  p.  307. 


134  COPPER  REFINING 

casting  wire  bars.  When  the  length  of  the  mold  is  in  line 
with  the  stream  of  copper  there  is  a  decided  wash,  which 
tends  to  make  " splashes7'  and  "cold  sets"  in  the  finished 
bar.  Then  in  order  to  bring  successive  molds  in  line  it  is 
necessary  to  start  and  stop  the  machine,  which  jars  the 
partially  set  castings  just  made  and  tends  to  make  "edges," 
while  the  Walker  wheel  runs  continuously,  the  pouring 
being  done  on  the  fly,  so  to  speak.  The  lost  motion  and 
jarring  of  the  link-belt  type  of  machine  was  the  chief  rea- 
son for  its  abandonment  in  addition  to  higher  maintenance 
costs.  The  Clark  machine  has  the  advantage  that  it  will 
accommodate  the  giant  wire  bars  supplied  to  a  certain  ex- 
tent to  foreign  wire  mills. 

As  the  size  of  furnace  began  to  outstrip  even  mechanical 
means  for  casting,  always  having  in  mind  the  desirability 
of  completing  a  furnace  cycle  in  an  even  twenty-four  hours, 
the  multiple-lip  ladle  was  devised.  In  this  way  as  many  as 
four  wire  bars  or  eighteen  ingots  are  poured  at  one  move- 
ment of  the  intercepting  ladle.  Then  came  the  twin  ladle, 
and  it  was  demonstrated  that  casting  could  be  conducted 
at  the  rate  of  over  100,000  Ib.  an  hour.  This  could  again  be 
doubled  by  using  two  wheels. 

The  bars  are  carefully  inspected  before  shipment  and 
any  minor  defects  such  as  small  "fins"  cut  off  with  a  ham- 
mer and  chisel.  The  temperature  of  the  bosh  water  and 
also  of  the  molds  has  much  to  do  with  the  appearance  of 
the  finished  product,  although  color  due  to  slight  films  of 
oxide  has  not  the  importance  attached  to  it  to-day  that  was 
formerly  the  case.  The  bars  should  be  dropped  into  the 
water  as  soon  as  possible  after  solidification  is  complete. 
The  molds,  which  upset  but  do  not  enter  the  water  and 
are  sprayed  to  cool  them,  should  retain  a  temperature  just 
sufficient  to  enable  them  thoroughly  to  dry  and  properly 
take  the  bone  ash  wash  which  is  applied  to  prevent  the 
bars  from  sticking.  On  the  other  hand,  the  temperature 
of  the  bosh  water  itself  must  be  held  down  by  a  circulating 
system  to  a  point  where  it  does  not  interfere  with  clear  vision 
due  to  clouds  of  steam  given  off. 


FURNACE  REFINING  135 

Various  materials  have  been  tried  out  as  substitutes  for 
the  rather  expensive  bone  ash  used  in  painting  molds.  It 
is  necessary  that  the  substance  used  should  have  a  high 
melting  point,  have  no  tendency  to  enter  into  chemical 
combination  with  molten  copper,  and  particularly  give  off 
no  gaseous  constituent  under  heat  which  would  tend  to  make 
the  casting  porous. 

Lampblack  serves  the  purpose,  but  is  not  readily  applied. 
A  smoked  mold  gives  a  beautiful  casting.  Pulverized 
silica  is  another  substitute,  but  as  the  finish  given  is  not  en- 
tirely satisfactory  its  use  is  confined  to  anode  work. 

Bone  ash  should  be  thoroughly  calcined  to  be  free  from 
grease,  which  evolves  gases  in  the  mold.  A  typical  analy- 
sis of  bone  ash  is : 

Ca3P2O8 97. 69  per  cent 

SiO2 1 . 01  per  cent 

MgO 1 . 02  per  cent 


Total 99 . 72  per  cent 

Many  of  the  minor  physical  defects  in  refined  copper  can 
be  traced  to  improper  mold  wash. 

Another  matter  which  needs  attention  is  the  chemical 
composition  of  the  water  used  for  cooling  the  copper.  It 
is  quite  possible  by  constant  evaporation  coupled  with  an 
impure  make-up  water  to  obtain  concentrations  of  sodium 
chloride  and  other  salts  which  will  crystallize  out  on  the 
surface  of  the  hot  molds  in  sufficient  quantity  to  cause  pin- 
holes  in  the  bars. 

The  molds  themselves  can  be  made  of  either  iron  or  cop- 
per, but  the  latter  gives  the  best  results.  For  many  years 
molds  were  made  by  filling  a  built-up  iron  box  with  molten 
copper  into  which  was  thrust  a  core  imprinting  the  desired 
shape.  This  method  involves  a  complicated  calculation 
of  shrinkages  starting  with  an  iron  mother  core  to  produce 
a  copper  mother  mold,  which  in  turn  makes  copper  cores 
from  which  the  final  molds  are  obtained.  To-day  this 
has  been  largely  superseded  by  molding  in  sand  with  an 
ordinary  wooden  pattern. 


136  COPPER  REFINING 

A  mold  generally  fails  by  burning  or  by  warping. 
Spongy  spots  can  be  hammered  up,  but  a  mold  should  be 
discarded  before  any  appreciable  sponginess  is  found  in  the 
castings  produced.  Cracks  in  a  mold  tend  to  hold  water 
which  steams  while  the  copper  is  being  poured. 

The  general  question  of  defects  in  refined  copper  and 
their  causes  will  be  discussed  more  in  detail  in  the  next 
chapter. 

The  Furnace  Cycle. — A  normal  cycle  for  a  large  modern 
furnace  would  be  about  as  follows: 

Charging 3  hours 

Melting  to  "flatness" 5  hours 

Recharging 1  hour 

Melting  to  "afloat" 5  hours 

Blowing 2  hours 

Skimming- 1  hour 

Poling 2  hours 

Pouring 5  hours 

24  hours 

The  fire  is  grated  and  rebuilt  during  the  charging  period. 

Slag  Formation. — The  amount  of  slag  made  naturally 
varies  with  the  nature  of  the  charge.  It  is  a  minimum  when 
melting  straight  cathodes.  The  addition  of  wire  mill  scrap 
at  once  increases  it.  In  the  anode  furnace  we  have  in  addi- 
tion to  blister  copper,  black  copper,  silver  building  slag, 
liberator  tank  residues,  cement  copper,  secondary  scrap 
copper  and  miscellaneous  clean-up  material. 

It  is  possible  to  melt  cathodes  in  a  basic  furnace  and  keep 
the  slag  made  down  to  a  small  fraction  of  1  per  cent  of  the 
weight  of  the  charge.  In  order  to  do  this  care  must  be 
taken  to  keep  clay  used  for  luting  up  the  furnace  doors 
from  getting  into  the  bath  and  the  fire  must  be  so  managed 
as  to  blow  as  little  coal  ash  over  as  possible  as  it  increases 
slag  by  uniting  with  copper  oxide. 

On  the  other  hand,  a  foul  anode  charge  may  make  20  per 
cent  of  slag  and  still  represent  good  work.  Representative 
slag  assays  are  given  in  Table  40. 


FURNACE  REFINING 

TABLE  40. — REPINING  FURNACE  SLAG  ASSAYS 


137 


Per  cent 

Copper 

Iron 

Nickel 

Insoluble 

Lime 

Sulphur 

Arsenic 

Anode  slag  
Wirebar  slag  

40.51 
48.55 

• 
5.30 
2.99 

6.99 

38.74 
41.73 

0.25 
0.53 

0.17 
0.19 

0.21 

When  nickel  is  present  in  quantity  an  infusible  scoria  is 
formed  which  carries  with  it  a  mass  of  metallics.  Tin 
likewise  forms  a  blanket  of  stannic  oxide,  which,  however, 
can  be  fluxed. 

Although  the  elimination  of  the  elements  occurs  in  the 
order  of  their  oxidizability  the  separations  are  not  sharp, 
and  while  a  certain  classification  of  impurities  for  further 
treatment  can  be  roughly  made,  the  losses  in  any  such  plan 
to  separate  lead,  tin,  zinc,  etc.,  must  necessarily  be  large, 
due  to  overlapping. 

In  the  same  way  the  copper  slagged  off  carries  with  it 
some  gold  and  silver.  When  anode  assays  over  a  period 
of  time  are  compared  with  corresponding  anode  slag  assays 
the  silver  appears  too  high.  This  is  due  to  the  large  amount 
of  silver  brought  into  the  charge  in  silver-building  slags, 
which  is  not  efficiently  collected  in  the  anodes  but  is  passed 
on  to  the  anode  slag  retreatment  stage  before  it  is  again 
associated  with  the  copper.  Table  41  gives  some  assays 
where  true  comparisons  can  be  made  for  both  anode  and 
wire  bar  furnaces. 

TABLE  41. — RELATIVE  SLAGGING  OF  GOLD  AND  SILVER 


Per  cent 
copper 

Oz.  per 

ton  silver 

Oz.  per 
ton  gold 

Anode  slag  

32  50 

14.3 

0.27 

Anodes  

99  44 

81  48 

2.299 

Compensated  slag  assav  

99.44 

43.75 

0.826 

Slagging  ratio  
Wirebar  slag. 

1.000 

48  55 

0.537 
1.35 

0.359 
0.0140 

Wirebars 

99  94 

1  40 

0  0104 

Compensated  slag  assav  

99.94 

2.76 

0.0286 

Slagging  ratio  

1.000 

1.97 

2.75 

138  COPPER  REFINING 

The  surprising  difference  in  the  behavior  of  the  wire  bar 
furnace  is  due  to  the  fact  that  the  silver  and  gold  in  the  cat- 
hodes is  not  alloyed  with  the  copper  but  mechanically  en- 
tangled as  wandering  anode  slimes.  Upon  fusion  some  of 
this  slime  floats  and  is  skimmed  off  with  the  slag.  It  is  not 
all  released,  however,  as  we  see  not  only  by  direct  assay  of 
the  wire  bars  but  from  the  lowering  of  electrical  conduc- 
tivity by  one  or  two  per  cent  when  wire  bars  are  compared 
with  cathodes. 

The  general  question  of  making  anodes  from  foul  material 
will  be  reserved  for  treatment  in  Chapter  X. 

Thermal  Efficiency  of  the  Refining  Furnace. — The  coal 
consumption  of  a  large  modern  refining  furnace  under 
average  operating  conditions  is  about  12  per  cent  of  the  good 
product  made.  As  the  coal  used  in  obtaining  this  figure 
runs  about  13,750  b.t.u.  per  pound  the  heat  requirement  is 
13,750  X  0.12  or  1650  b.t.u.  per  pound  of  copper  produced. 

If  we  take  the  mean  specific  heat  of  copper  at  0.11  and 
the  latent  heat  of  fusion  as  80  b.t.u.  per  Ib.  and  assume  that 
copper  is  charged  at  70°F.  and  poured  at  2170°F.  we  have 
the  net  heat  input  required  0.11  X  2100  +  80  or  311  b.t.u. 
per  Ib.,  giving  an  efficiency  of  311  -r-  1650  or  about  19  per 
cent.  Waste  heat  boilers  now  recover  heat  equivalent  to 
at  least  6  Ib.  water  from  and  at  212°F.  per  Ib.  coal  burned 
in  the  furnace.  This  is  equivalent  to  966  X  6  X  0.12  or 
696  b.t.u.  per  Ib.  of  copper,  bringing  the  overall  thermal 
efficiency  up  to  1007  -r-  1650  or  about  61  per  cent. 

It  will  therefore  be  seen  that  operating  as  a  simple 
reverberatory  the  furnace  is  very  wasteful  but  that  when 
there  is  a  demand  for  steam  the  combined  furnace  and 
boiler  is  nearly  as  economical  of  heat  as  would  be  an  electric 
furnace.  The  general  question  of  waste  heat  boilers  will 
be  dealt  with  in  Chapter  XI. 

Retreatment  of  Slags. — Slags  made  from  normal  anode 
and  refining  furnace  charges  require  reduction  with  the 
substitution  of  another  base  to  free  the  copper.  As  the 
slags  from  cathode  charges  carry  very  small  values  in  gold 
and  silver  they  are  sometimes  kept  separate  from  the  anode 


FURNACE  REFINING  139 

slags  and  the  copper  recovered  therefrom  made  into  casting 
copper  in  order  to  avoid  the  cost  of  passing  this  metal  through 
the  anode  furnace  and  tank  house  steps  of  the  process. 

The  simplest  means  of  reduction  is  a  small  blast  furnace 
operated  with  limestone  and  pyrites  cinder  as  fluxes  and 
making  black  copper  as  a  product.  A  furnace  much  smaller 
than  44  in.  X  66  in.  gives  trouble  in  keeping  the  crucible 
open  and  as  even  this  small  size  has  a  capacity  in  excess 
of  the  requirements  of  most  refineries,  it  is  often  necessary 
to  run  intermittent  campaigns. 

If  true  black  copper  is  made— say  97  per  cent  copper 
carrying  considerable  iron — a  very  bad  slag  is  inevitable, 
the  copper  running  2  or  3  per  cent  therein.  As  a  rule  there 
is  enough  sulphur  in  the  iron  flux  used  to  produce  a  little 
matte,  and  a  copper  running  about  94  per  cent  and  a  slag 
of  0.6  per  cent  is  made.  This  product  delays  the  work  in 
the  anode  furnace  unless  it  is  systematically  charged  back 
day  by  day,  as  small  quantities  of  sulphur  take  some  time  to 
blow  out  of  an  anode  charge. 

Where  a  smelter  is  operated  in  conjunction  with  a  refin- 
ery, these  furnace  slags  are  usually  added  to  the  ore  charge. 
The  reducing  action  in  a  matting  furnace  is  hardly  strong 
enough  thoroughly  to  decompose  these  high  grade  silicates 
and  a  noticeable  increase  in  oxidized  slag  losses  always 
follows  their  addition  to  the  charge. 

They  are  also  very  difficult  accurately  to  sample  on 
account  of  the  metallics  contained  and  the  refinery  will 
suffer  and  the  smelter  correspondingly  gain  unless  special 
supervision  is  exercised  over  this  sampling,  as  the  workmen 
naturally  avoid  metallics  in  taking  the  sample  as  they  give 
such  trouble  in  the  crushing. 

Finally  nothing  is  gained  in  cost  by  matting  these  slags, 
as  the  advantage  of  the  larger  scale  of  operation  and  some- 
what lower  slag  losses  is  offset  by  the  sampling  costs  and 
complication  of  adding  sulphur  to  this  high  grade  sulphur- 
free  material  only  to  blow  it  out  again  in  the  succeeding 
converting  operation.  The  anode  furnace  gains  somewhat 
however  in  not  having  any  low  grade  black  copper  to  treat. 


140  COPPER  REFINING 

Future  Developments. — Great  progress  has  been  made  in 
the  development  of  the  refining  furnace  and  not  much 
remains  to  be  done  along  purely  structural  and  mechanical 
lines.  In  twenty  years  the  24-hour  charge  capacity  has 
been  changed  from  20  tons  to  250  tons  and  furnaces  of 
greater  capacity  could  easily  be  built  were  there  any  justifi- 
cation for  it.  The  copper  is  handled  mechanically  both  in 
and  out  of  the  furnace.  The  hearth  is  now  constructed  of 
magnesite  brick  when  a  corrosive  charge  is  to  be  handled. 
The  walls  and  verb  and  sometimes  the  entire  roof  are  now 
built  of  chrome  or  magnesite  brick  to  withstand  the  copper 
oxide  wash. 

In  firing  the  furnace  there  is  yet  some  room  for  progress, 
but  when  it  is  considered  that  an  open  hearth  steel  furnace 
takes  four  times  as  much  fuel  per  ton  of  product,  it  will  be 
seen  that  there  is  not  the  same  justification  for  expensive 
gas  firing  equipment.  Either  oil  or  gas  firing  may  be 
successfully  applied,  however,  and  commercial  factors 
may  some  day  call  for  their  introduction.  The  poling 
operation  as  already  pointed  out  will  doubtless  in  time  be 
replaced  by  gas  reduction. 

Metallurgically  the  anode  operation  is  quite  satisfactory. 
The  melting  of  cathodes  is  not  in  such  good  shape,  how- 
ever, as  here  what  should  be  a  simple  melting  is  expanded 
into  a  complete  refining.  This  has  been  attacked  in  two 
ways — by  the  use  of  the  electric  furnace  and  by  continuous 
melting. 

The  electric  furnace  is  not  very  promising  as  it  has 
already  been  shown  that  with  a  waste  heat  boiler  attached 
the  heat  efficiency  of  the  reverberatory  is  high  enough  to 
give  electric  heat  from  steam  driven  generators  a  hopeless 
handicap.  Then  the  high  electrical  conductivity  of  copper, 
the  shape  of  the  cathodes  to  be  melted  and  the  necessity 
of  having  the  product  absolutely  free  from  bits  of  charcoal 
or  other  resistor  material,  introduce  limitations  in  the  design 
of  the  furnace. 

It  must  also  be  remembered  that  there  is  a  slight  re- 
fining done  in  melting  cathodes  as  shown  by  the  concen- 


FURNACE  REFINING  141 

tration  of  impurities  in  the  slag  as  well  as  the  effective 
elimination  of  sulphur  brought  in  as  entrained  sulphates 
from  the  electrolyte. 

Continuous  melting  is  more  attractive.  The  method 
already  described  of  charging  large  quantities  of  cathodes 
during  pouring  and  carefully  excluding  all  acid  material 
from  contact  with  the  bath  to  avoid  the  formation  of  slag 
has  been  for  some  years  in  successful  use  and  might  be 
called  semi-continuous  melting.  It  is  obvious  that  the 
principle  could  be  extended  until  the  amount  charged 
just  balanced  the  amount  poured,  but  this  requires  slow 
casting  rates  in  order  to  keep  the  heat  up  and  introduces 
some  practical  difficulties.  The  sulphur  is  controlled  in 
this  method  by  inserting  piles  of  cathodes  which  dg,  not 
fully  submerge  in  the  molten  bath.  This  gives  the  oppor- 
tunity for  a  little  oxidation  of  the  exposed  sheets  before 
the  mass  softens  and  becomes  immersed. 

Various  patents  have  been  issued  for  plans  which  virtu- 
ally combine  a  blast  furnace  melting  with  a  reverberatory 
fore-hearth.  Copper  is  so  readily  contaminated  by  con- 
tact with  products  of  combustion,  etc.,  however  that  it  is 
difficult  to  apply  methods  of  direct  melting  without  having 
to  go  through  a  large  part  of  the  regular  furnace  treatment 
afterwards. 

Perhaps  the  weakest  point  in  the  present  reverberatory 
practice  lies  hi  the  system  of  utilizing  labor.  Custom  has 
placed  a  crew  in  charge  of  each  furnace.  Originally  a 
night  crew  did  the  melting  and  rabbling  and  some  of  the 
poling  while  the  day  crew  would  attend  to  the  ladling  and 
charging.  As  a  number  of  men  were  required  for  hand 
ladling  this  fitted  in  with  the  heavy  work  of  charging.  All 
the  furnaces  were  supposed  to  be  ready  to  cast  early  in 
the  morning  and  the  crew  kept  hours  depending  upon  the 
furnace  which  would  generally  get  progressively  later 
during  the  week,  the  over  Sunday  period  being  used  to 
catch  up. 

When  mechanical  ladling  was  introduced  it  became 
necessary  to  organize  special  labor  gangs  to  assist  in  the 


142  COPPER  REFINING 

charging  of  the  furnaces  which  became  much  larger  and 
with  a  reduced  ladling  crew. 

Mechanical  charging,  however,  eliminated  this  trouble 
and  the  necessary  crew  for  handling  a  furnace  is  very  small. 
The  old  system  is  therefore  beginning  to  give  way  to  one 
where  instead  of  operating  in  unison  the  furnaces  are  in- 
tentionally staggered.  In  this  way  they  may  be  skimmed 
one  after  another,  for  example,  and  a  casting  crew  can  be 
organized  who  will  go  from  one  furnace  to  another  and  be 
replaced  by  a  new  shift  at  a  predetermined  hour,  making 
the  men  entirely  independent  of  the  behavior  of  individual 
furnaces.  When  this  is  fully  accomplished  furnace  work 
will  become  a  less  arduous  and  less  specialized  occupation. 


CHAPTER  IX 
THE  REQUIREMENTS  OF  REFINED  COPPER 

In  order  to  satisfy  market  requirements,  prime  electroly- 
tic copper  wirebars  and  cakes  must  possess  high  electrical 
conductivity,  sufficient  ductility  and  physical  soundness. 
From  the  refiner's  point  of  view  the  product  must  also  be 
free  from  commercial  quantities  of  precious  metals.  Ingots 
are  given  a  little  more  leeway  in  conductivity  and  relieved 
of  restrictions  as  to  soundness,  inasmuch  as  they  are  in- 
tended for  remelting. 

This  general  subject  may  advantageously  be  discussed 
under  the  headings  of  conductivity,  pitch,  ductility,  castings 
and  dimensions. 

Electrical  Conductivity. — It  has  come  to  be  generally 
assumed  that  if  a  wirebar  is  a  sound  casting  and  a  sample  of 
the  copper  drawn  into  wire  shows  a  satisfactory  conduc- 
tivity, the  product  may  be  accepted  as  prime  electrolytic 
(or  Lake)  copper.  This  assumption  is  not  strictly  true,  as 
will  be  shown  later,  but  inasmuch  as  the  preparation  of  the 
sample  of  wire  for  testing  in  itself  demonstrates  within 
certain  limits  the  ductility  of  the  copper,  as  the  test  may 
be  easily  and  quickly  made  and  as  no  other  test  appears  to 
be  of  general  application,  conductivity  will  probably  remain 
as  the  chief  criterion  of  commercial  copper. 

Impurities  in  copper  may  be  divided  into  three  groups  as 
regards  their  effect  upon  conductivity:  Those  which  are 
insoluble  in  copper,  those  which  are  partially  soluble  and 
those  which  are  completely  soluble — remembering  that  we 
are  interested  only  in  mixtures  where  copper  enormously 
predominates. 

The  elements  which  are  insoluble  can  evidently  have 
but  little  effect  upon  the  conductivity,  inasmuch  as  the 

143 


144 


COPPER  REFINING 


total  quantity  of  any  one  impurity  is  always  far  below  0.1 
per  cent  and  the  impurity  itself  has  a  certain  specific 
conductivity. 

The  class  which  is  partially  soluble  exerts  a  very  marked 
depression  on  the  conductivity,  as  such  elements  form  low 
melting  point  alloys  with  copper  which  crystallize  out  as 
eutectics,  forming  a  matrix  surrounding  crystals  of  copper 
like  the  mortar  joints  of  a  mosaic.  This  matrix  may  be 
quite  bulky,  as  part  of  the  copper  is  added  to  the  quantity 
of  impurity  present. 


0.02 


o.io 


O.U4  O.u6  0.08 

Percent  Impurities 
FIG.  27. — Effect  of  impurities  on  conductivity. 


0.12 


The  few  elements,  such  as  silver,  that  form  solid  solutions 
with  copper  have  a  relatively  mild  effect  upon  the  conduc- 
tivity, which  is  quite  negligible  in  view  of  the  small  quanti- 
ties present. 

About  fifteen  years  ago  the  writer  conducted  a  systematic 
investigation  into  the  relation  between  impurities  and 
copper.1  Each  element  in  question  was  added  in  varying 
amounts  to  crucible  melts  of  high-grade  copper  wire  and  the 
conductivity  of  an  annealed  sample  of  wire  drawn  from 
the  cast  compared  with  the  assay  of  the  same  wire  for  the 
impurity  added.  The  results  of  this  work  are  grouped 
together  in  Fig.  27. 

1  Trans.,  A.  I.  M.  E.,  vol.  xxxvi,  p.  18. 


REQUIREMENTS  OF  REFINED  COPPER          145 


It  will  be  noted  at  once  that  the  elements  which  are 
known  to  make  copper  brittle,  such  as  lead,  bismuth  and 
tellurium,  have  but  slight  effect  upon  the  conductivity, 
while  those  which  make  excellent  bronzes,  such  as  phos- 
phorus, aluminum  and  silicon,  have  caused  a  marked  de- 
pression. It  is  difficult,  however,  to  make  any  practical 


62  100.4 

60  100.2 

68  100.0 

56  99.8 

54  99.6 

52  99.4 


Per  Cent,  Reduction  in  Area 
10      20      30      40      50      60     70       80     90     100 


1 

48   £ 

99.0 

H 

.> 

£ 

46    o 

96.8 

I 

-1 

98.6 

w 

42 

98.4 

i± 
H 

40 

98.2 

38 

96.0 

86 

97.8 

34 

97.6 

32 

97.4 

30        97.2 


A 

Per  Cent   Conductivity 
Matthiessen's  Standard 

/ 

V 

y 

/ 

c 

\ 

/ 

/ 

\ 

\ 

/ 

/ 

• 

\ 

/ 

/  Tensile  Strength 
'  ^In  Ib.per  sq.in- 

\ 

f 

0 

r  onj 

jinal 

Area 

/ 

\ 

/ 

/ 

\ 

V 

/ 

x, 

X 

/ 

X 

*s 

/ 

N. 

I 

FIG.  28. — Annealed  copper  rod  of  different  sizes  drawn  to  No.  12  B.  &  S. 

application  of  these  results,  because  the  presence  of  one 
impurity  often  neutralizes  the  effect  of  another,  and  copper 
which  would  be  expected  to  be  of  indifferent  quality  based 
upon  these  figures  may  prove  to  be  excellent.  This  is  the 
reason  that  chemical  analysis,  apart  from  the  difficulties 
attending  the  accurate  determination  of  traces  of  impuri- 
ties, has  proved  of  little  value  in  judging  copper  as  com- 

10 


146 


COPPER  REFINING 


pared  with  the  more  practical  test  of  forging  and  drawing 
followed  by  a  measurement  of  conductivity. 

By  common  consent  conductivity  has  come  to  be  ex- 
pressed in  percentage  of  the  value  of  0.141729  international 
ohm  at  0°C.  found  by  Matthiessen  for  a  meter-gram  of 
supposedly  pure  copper  over  fifty  years  ago.  The  speci- 
fications of  the  American  Society  for  Testing  Materials 
have  established  a  minimum  requirement  equivalent  to  98.5 
per  cent  for  wirebars  and  cakes  and  97.5  per  cent  for  ingots, 


67     100.6 

CK        100  4  i 

)          20         40         60          8( 

Amperes 
)         100        120       140       160         180       200      220 

t 

63      100.2 
61      100.0 
59        99.8 
c  57    ^99.6 
3  55  ?  99.4 
H  53  |  99.2 
£51  1  99.0 
g49   j  98.8 
£  47  g  98.6 
~  45  £  98.4 

V>              Pi 

£  43        98.2 
41        98.0 
39        97.8 

It 

nsi 

eS 

ren 

gth. 

>^ 

_ 

\ 

f 

> 

X 

\l 

Pe 

r  Cent 

\ 

\\ 

Con 

ductivit 

\ 

\ 

^ 

\ 

\ 

\ 

\ 

I 

\ 

\ 

I 

) 

m 

g 

O> 

3 

3 

] 

\ 

• 

/ 

/    i 

W 

«2 

"3 
p 

w 

\ 

I 

3 
M 

^^^ 

p 

T, 

!S6_] 

ier_ 

5q.-7 

n-o 

iOi 

ffi-7 

•T]-I\ 

\ 

37       97.6 

35          Q7  A 

1 

^T^ 

— 

1    el 


il 

•cw 


FIG.  29. — Electric  annealing  of  No.  12  B.  &  S.  copper  wire. 

determinations  being  made  upon  annealed  samples.  Aver- 
age copper  on  the  market  runs  considerably  higher  than 
these  figures,  much  of  it  being  over  100  per  cent  as  com- 
pared with  the  imperfectly  purified  standard  copper  of 
Matthiessen.  The  purest  cathode  copper — that  is,  an 
annealed  sample  of  wire  drawn  directly  from  a  cathode 
without  melting — may  run  over  102  per  cent;  it  is  unusual, 
however,  for  copper  that  has  been  melted  to  run  over  101 
per  cent.  The  lowering  of  conductivity  by  melting  is  due 


REQUIREMENTS  OF  REFINED  COPPER 


147 


chiefly  to  the  fact  that  much  of  the  impurity  content  of  the 
cathode  is  present  merely  as  mechanical  contamination,  a 
certain  proportion  of  which  a  melting  will  incorporate 
chemically.  Then  there  is  always  the  opportunity  for  the 
copper  to  absorb  impurities  from  outside  sources  during 
furnace  refining. 


97.0 


97.5 


Per  Cent,  Conductivity 
96.0          98.5  99.0 


99.5         100.0        100.5 


32 

FIG.  30. — Relation  between  conductivity  and  tensile  strength  in  copper  wire. 

In  drawing  copper  into  wire  the  coarse  crystalline  struc- 
ture of  cast  copper  is  broken  down  into  a  fibrous  structure 
much  harder  and  stronger  but  of  a  lower  electrical  con- 
ductivity, due  perhaps  to  a  rearrangement  of  the  system  of 
series-parallel  circuits  made  by  copper  crystals  and  matrix. 
The  depression  of  conductivity  bears  a  direct  relation  to 
the  degree  of  hardness  attained,  as  shown  in  Figs.  28,  29  and 


148  COPPER  REFINING 

30. l  It  will  be  seen  from  Figs.  29  and  30  that  the  softening 
of  a 'wire  by  annealing  has  a  progressive  character  in  the 
exact  reverse  direction  from  the  hardening  by  working  the 
metal. 

In  order  to  arrive  at  a  definite  standard  for  comparing 
conductivities,  therefore,  wire  is  annealed  in  order  to  ob- 
literate the  somewhat  uncertain  effect  of  hard  drawing 
before  a  determination  is  made.  On  the  other  hand,  either 
a  partial  annealing  or  an  overheating  or  " burning"  of 
the  wire  will  give  results  below  the  truth.  It  is  therefore 
necessary  to  bring  the  wire  to  incipient  red  heat  (about 
900°F.,  or  480°C.)  for  a  brief  time  in  order  to  get  consistent 
results.  All  errors,  however,  are  on  the  safe  side,  as  it 
is  impossible  to  get  unduly  high  conductivity  by  faulty 
manipulation. 

Pitch  or  Set — The  pitch,  or  set,  of  copper  is  an  expres- 
sion used  to  define  the  appearance  of  the  free  surface  of  a 
cast  copper  bar;  that  is,  the  bar  may  be  of  high,  normal  or 
low  pitch  or  it  may  be  "in  set"  or  "out  of  set." 

Molten  copper  readily  dissolves  the  "various  gases 
with  which  it  is  brought  in  contact  in  the  refining  furnace, 
such  as  oxygen,  carbon  monoxide  and  dioxide,  and  sul- 
phur dioxide.  As  the  copper  cools  and  sets  in  the  mold 
excess  gas  is  given  off,  and  the  proper  control  of  this 
phenomenon  is  the  secret  of  good  furnace  refining.  As 
has  been  outlined  in  the  last  chapter,  the  bath  after  melting 
is  allowed  to  oxidize  and  the  resulting  cuprous  oxide  dis- 
solves and  acts  as  a  scorifying  agent,  the  scoria  is  skimmed 
off  and  the  excess  cuprous  oxide  reduced  by  poling.  A 
button  of  copper  cast  in  a  "say-ladle"  gives  a  characteristic 
set  surface  and  fracture  for  each  stage,  the  finished  product 
showing  a  slightly  swelling  surface  without  deep  wrinkles 
and  a  rose-red  silken  fracture.  Should  the  poling  not  be 
carried  far  enough,  the  set  surface  will  show  a  depression 
and  the  metal  will  be  hard  and  somewhat  sonorous.  Met- 
allographic  examination  will  show  thick  veins  of  cuprous 

1  See  Addicks,  Trans.,  American  Institute  of  Electrical  Engineers,  vol. 
22,  p.  695. 


REQUIREMENTS  OF  REFINED  COPPER          149 


150  COPPER  REFINING 

oxide  eutectic,  and  careful  inspection  of  the  depressed  set 
will  reveal  "niggerheads,"  or  minute  entrances  to  con- 
siderable cavities  within.  Figure  31  shows  a  low  set  bar 
which  has  been  planed  in  longitudinal  cross-section  and 
one  or  two  of  these  cavities  are  plainly  shown.  Such  a 
bar  may  show  red  shortness  in  rolling  or  merely  a  slight 
hardness,  but  the  cavities  will  roll  out  into  small  "cold  sets " 
and  are  probably  the  main  cause  of  splinters  and  slivers 
frequently  met  with  in  wire  drawing.  This  phenomenon 
is  identical  with  "  piping "  in  steel  and  brass,  which  are 
always  cast  on  end,  the  set  surface  being  sheared  off  and 
rejected.  Casting  of  copper  on  end  has  been  advocated 
but  never  generally  adopted  except  for  wedge  cakes  used 
for  purposes  where  a  special  mirror  polish  is  required  in  the 
rolled  article  and  certain  special  shapes  such  as  round 
billets. 

If,  on  the  other  hand,  poling  is  pushed  too  far,  a  reverse 
condition  sets  in.  The  surface  bulges  on  setting  and  finally 
"  spews  over,"  throwing  out  worm-like  excrescences.1  This 
is  analogous  to  the  well-known  "  spitting "  of  silver.  Such 
a  bar  shows  perfect  softness  and  malleability  in  rolling, 
but  the  worms  have  to  be  chipped  off,  and  overpoled 
copper  is  not  a  satisfactory  product.  In  general,  however, 
the  best  pitch  is  one  as  high  as  consistent  with  the  avoidance 
of  overpoling. 

Overpoled  copper  shows  a  slight  depression  in  con- 
ductivity, as  may  be  seen  from  the  curve  marked  oxy- 
gen in  Fig.  27.  The  microscope  shows  the  formation  of 
some  compound,  presumably  involving  the  gases  of  re- 
duction or  possibly  sulphur.  The  latter  element  is  pres- 
ent as  included  electrolyte  in  the  cathode  and  in  the 
products  of  furnace  combustion,  and  until  the  sulphur 
has  been  worked  out  a  charge  of  blister  copper  in  an  anode 
furnace  will  give  a  test  button  which  will  "throw  a  worm" 
quite  similar  to  that  of  overpoled  copper.  Cases  of  myster- 
ious aging  and  season  cracking  in  brass  have  been  traced 
to  traces  of  sulphur,  and  it  must  be  remembered  that  Lake 

1  Hofman,  Hayden  and  Hallowell,  Trans.,  A.  I.  M.  E.,  vol.  38,  p.  175. 


REQUIREMENTS  OF  REFINED  COPPER          151 

copper,  which  had  at  one  time  a  special  reputation  in  the 
manufacture  of  cartridge  brass,  did  not  enter  the  refining 
furnace  with  the  possibility  of  faulty  separation  from  a 
sulphate  electrolyte  in  its  past  history. 

An  overpoled  charge  of  copper  exhibits  instability  to 
pitch,  and  the  mere  exposure  to  the  air  long  enough  to 
restore  its  depleted  oxygen  content  is  not  sufficient  to 
stabilize  this  wabbly  pitch.  It  is,  therefore,  necessary 
to  rescorify  the  entire  charge,  another  indication  that  a  new 
compound  is  formed  in  overpoling.  Curiously  enough, 
the  addition  of  a  minute  proportion  of  metallic  lead  to  an 
overpoled  charge  will  stabilize  it,  but  there  is  great  danger 
of  making  the  copper  brittle.  This  corrective  influence 
might  also  point  to  the  presence  of  sulphur,  inasmuch  as 
while  lead  is  insoluble  in  copper,  leady  copper  matte  is 
readily  formed,  and  under  special  conditions  lead  and 
copper  can  be  made  to  unite  as  they  do  in  Allan's  metal  and 
in  many  Oriental  bronzes. 

The  reason  that  foundries  have  such  difficulty  in  making 
sound,  high-conductivity  castings  of  pure  copper  is  that 
they  get  the  copper  badly  below  pitch  in  melting,  so  that 
they  find  it  necessary  to  add  various  deoxidizing  agents. 
Much  phosphor  bronze  is  nearly  pure  copper,  the  phos- 
phorus having  vanished  as  volatile  phosphorus  pentoxide 
and  merely  served  as  a  substitute  for  poling. 

Copper  which  contains  a  considerable  proportion  of  some 
other  element — as  in  the  case  of  arsenical  Lake  running 
0.5  per  cent  arsenic — will  show  a  much  coarser  wrinkling 
of  the  set  surface  and  to  the  practiced  eye  has  quite  a 
different  appearance  from  pure  copper.  In  the  same  way 
copper  which  has  been  poled  with  oil  presents  a  peculiar 
appearance.  Low  set  copper  has  heavy  wrinkles  and  may 
cause  cold  sets  in  rolling. 

Bars  which  break  in  the  rolls  are  frequently  found  to 
possess  a  large  cavity  near  the  upper  center  as  shown  in 
Fig.  32,  due  to  a  lowness  of  set.  A  perfect  set  should  be 
high  and  rounding  like  a  mercury  meniscus  and  not  show 


152 


COPPER  REFINING 


a  low  level  center  and  a  high  edge  close  to  the  sides  where 
the  mold  chilled  the  copper.  The  wrinkles  should  be  fine 
and  the  surface  solid. 

The  set  surface  carries  somewhat  more  dissolved  oxide 
than  the  body  of  the  bar.  In  Europe  it  was  at  one  time 
the  custom  to  plane  this  surface  off  before  rolling  when 
difficult  specifications  had  to  be  met. 


• 


FIG.  32. — Gas  cavity  in  low  set  bar. 

Ductility. — While  it  is  known  that  the  addition  of  exceed- 
ingly small  quantities  of  certain  elements,  such  as  lead, 
bismuth  or  tellurium,  will  make  copper  so  red-short  as  to 
fall  to  pieces  in  rolling,  it  is  quite  impossible  to  state  per- 
missible limits  of  such  elements  on  account  of  the  neutraliz- 
ing effects  or  various  other  impurities  which  may  or  may  not 
be  present.  In  general  it  is  safe  to  say  that  nearly  all 
troubles  from  brittleness  may  be  traced  to  low  pitch. 

Well-refined  high-grade  copper  can  be  hot  rolled  into 
quarter-inch  rod  which  after  pickling  free  of  scale  can  be 
cold  drawn  into  the  finest  wire  without  any  intermediate 
annealing.  In  practice,  however,  there  is  always  a  certain 
percentage  of  " breaks"  in  the  wire-drawing  machines,  and 
this  percentage  has  been  found  to  bear  a  direct  relation 


REQUIREMENTS  OF  REFINED  COPPER          153 

to  the  copper  contents  of  the  lot  of  wirebars  under  test. 
It  is,  therefore,  customary  to  impose,  in  addition  to  the 
conductivity  specification,  a  requirement  that  the  copper 
content  (including  silver)  shall  not  be  less  than  99.88  per 
cent,  a  figure  intended  to  represent  a  limit  of  99.90  per 
cent  after  allowing  a  margin  for  assay  precision. 

The  determination  of  traces  of  impurities  in  copper 
requires  long  experience.  This  is  particularly  true  in  the 
case  of  sulphur,  where  even  the  atmosphere  of  the  labora- 
tory must  be  free  from  sulphur  compounds. 

Mechanical  tests  on  samples  of  wire  as  checks  upon  the 
character  of  the  raw  copper  have  been  abandoned  because 
the  fabrication  of  the  wire  introduces  so  many  variables. 
The  former  tests,  as  a  rule,  consisted  of  measurements  of 
tensile  strength  and  torsion. 

Tensile  Strength  and  Torsion. — The  tensile  strength  of 
copper  wire  depends  upon  the  percentage  of  reduction  in 
area  after  annealing,  the  speed  and  the  temperature  of 
drawing.  With  proper  regard  to  these  factors,  almost  any 
copper  not  too  brittle  to  draw  can  be  made  to  fulfill  stand- 
ard telephone  wire  specifications  as  to  strength. 

The  usual  torsion  test  was  to  determine  the  number 
of  twists  a  No.  12  B.  &  S.  (2.05  mm.  diameter)  hard-drawn 
wire  would  stand  before  snapping.  This  test  is  perfectly 
worthless,  because  it  depends  entirely  upon  the  properties 
of  the  thin  skin  which  forms  on  wire  in  drawing.  A  wire 
as  it  comes  from  the  drawbench  when  put  in  the  testing 
machine  will  twist  uniformly  throughout  its  length  for  a 
few  turns  and  then  twist  tightly  at  one  point  and  snap. 
If  the  wire  is  scoured  with  emery  paper  or  dipped  for  a 
moment  in  nitric  acid  before  twisting,  the  skin  is  broken 
down  and  before  snapping  the  entire  length  will  coil  tightly, 
giving  several  times  the  number  of  total  twists  formerly 
obtainable.  This  skin  is  at  least  in  part  due  to  imperfect 
pickling  after  rolling,  as  by  the  use  of  a  nitric  acid  instead 
of  sulphuric  acid  pickle  on  the  rod  the  number  of  twists 
that  the  wire  will  stand  is  greatly  increased.  Particles  of 
oxide  scale  are  probably  rolled  into  the  metal,  and  sulphuric 


154  COPPER  REFINING 

acid,  while  it  will  dissolve  oxide,  cannot  cut  metallic  copper 
to  reach  it. 

Sometimes  bars  are  badly  overheated  in  the  heating 
furnace  before  rolling,  and  this  probably  causes  absorption 
of  gases  and  consequent  brittleness.  Altogether,  mechani- 
cal tests  have  been  entirely  dropped  as  a  check  on  the 
quality  of  the  copper  as  distinguished  from  the  finished 
wire. 

Castings. — In  order  to  get  a  perfect  casting  from  a  physi- 
cal point  of  view,  the  first  requisite  is  that  the  copper  should 
be  in  correct  pitch,  a  subject  which  has  already  been  dis- 
cussed at  some  length.  Small  adjustments  in  pitch  are 
made  by  adding  or  removing  charcoal  at  the  casting  ma- 
chine ladle.  Next,  the  temperature  of  the  copper  must  be 
high  enough  to  insure  good  fluidity  without  burning  the 
mold.  The  mold  itself  must  be  warm — that  is,  slightly 
above  the  boiling  point  of  water — in  order  to  avoid  any 
condensation  of  moisture  or  too  sudden  chilling  of  the 
copper  first  striking  it.  There  must  be  a  fine  film  of  a 
suitable  mold  wash  to  prevent  sticking  and  the  material 
used  must  be  free  of  any  volatile  matter.  The  mold  must 
be  solid  and  free  of  pores  which  would  entrap  water;  other- 
wise a  spongy  casting  would  result.  The  copper  must  flow 
in  quickly  but  evenly  and  the  casting  machine  must 
operate  so  smoothly  that  a  partially  set  bar  will  be  subject 
to  no  vibration  or  jars.  The  bars  must  be  quenched  in 
water  soon  after  setting  and  this  water  must  be  warm. 
Finally,  the  molds  must  be  free  of  warp  and  set  truly  level 
and  filled  to  a  given  mark  without  any  attempt  at 
adjustment. 

These  many  small  points  require  very  careful  supervision 
at  the  casting  wheel  and  carelessness  in  any  one  detail  will 
give  more  or  less  defective  product.  Bars  may  be  out  of 
size,  out  of  set,  porous  or  contain  splasjies,  fins,  cold  sets  or 
"fish." 

Out-of-size  bars  result  from  out-of-level  or  warped  molds, 
or  from  careless  filling  to  a  wrong  level.  Too  large  a  bar 
may  result  in  the  production  of  fins  in  the  rolls,  due  to 


REQUIREMENTS  OF  REFINED  COPPER          155 

crowding.  Molds  suspended  at  the  ends  tend  to  drop  in 
the  middle  and  pinch  at  the  top  center. 

Porosity  may  come  from  low  set  or  may  be  due  to  mois- 
ture. It  generally  follows  cold  or  porous  molds  or  too 
high  a  casting  temperature.  In  this  connection  it  may  be 
stated  that  copper  molds  turn  out  superior  castings  to  iron 
ones,  although  the  latter  are  more  or  less  in  use. 

Splashes  occur  when  the  copper  first  strikes  the  mold, 
being  thin  sheets  of  metal  which  splash  on  the  sides  and 
freeze.  Unless  they  are  knocked  off  with  a  suitable  tool 
before  they  are  submerged,  they  are  likely  to  leave  a  cold 
set  or  un welded  lamination  in  the  bar. 

Fins  are  sharp  high  edges  due  to  filling  the  mold  too 
rapidly  or  to  vibration  of  the  casting  machine  giving  a  wash 
to  the  bar  before  setting.  They  must  be  chiseled  off  the 
finished  bar  or  they  will  be  rolled  in  later. 

Cold  sets  occur  in  various  ways  whenever  molten  cop- 
per is  allowed  to  enfold  cold  splashes,  etc.  They  must  be 
chiseled  out  or,  if  extensive,  the  bar  rejected. 

Fish  are  bits  of  charcoal,  nails  or  other  foreign  bodies 
allowed  to  become  incorporated  in  the  casting.  They 
are  supposed  to  be  removed  by  a  "fisher"  during  pouring 
much  as  a  fly  is  rescued  from  a  glass  of  milk. 

The  color  of  cast  copper  depends  chiefly  upon  the 
time  allowed  to  elapse  between  pouring  and  quenching 
and  the  temperature  of  the  bosh  water.  The  beautiful 
ruby  shades  are  due  to  thin  films  of  the  lower  oxides,  and  a 
slight  change  from  the  proper  conditions  will  make  bars 
either  brassy  or  black.  Twenty  years  ago  good  color  was 
given  great  importance  at  the  rolling  mill,  but  it  has  nothing 
to  do  with  the  quality. 

Dimensions. — Each  refinery  issues  a  metal  schedule 
giving  the  weights  and  dimensions  of  the  market  shapes  it 
produces.  This  covers  wirebars,  cakes,  ingots,  ingot  bars 
and  billets.  Very  large  wirebars,  very  heavy  cakes  and 
billets  cast  on  end  are  poured  by  hand  from  a  crane 
ladle  and  the  additonal  cost  over  machine  casting  is 
reflected  in  a  premium  charged.  Ingotbars  are  merely 


156 


COPPER  REFINING 


three  ingots  cast  end  to  end  in  order  to  bring  the  weight 
up  to  a  figure  where  steamships  will  accept  them  without 
demanding  that  they  be  barreled. 

No  rigid  requirement  as  to  regularity  of  size  and  weight 
is  required  of  ingots  and  ingotbars  which  are  to  be  remelted. 
In  the  case  of  wirebars,  however,  the  American  Society 
for  Testing  Materials  specifications  give  a  tolerance  of 
5  per  cent  in  weight  and  y±  in.  in  any  dimension  except  that 
wirebars  may  vary  1  per  cent  in  length  and  cakes  3  per 
cent  in  any  dimension  greater  than  8  in. 

For  wirebars,  each  refinery  has  its  own  design  and 
proportions  of  point.  Different  rolling-mill  men  have 
such  contradictory  opinions  regarding  the  relative  advan- 
tages of  slight  changes  in  these  matters  that  it  is  probable 
that  one  design  is  about  as  good  as  another.  A  long  point 
enters  the  roll  more  easily  at  the  first  pass,  but  is  apt  to 
leave  more  scrap.  A  square  cross-section  is  desirable, 
because  a  flat  bar  is  apt  to  turn  over  in  the  rolls  when  part 
way  through.  The  first  pass  is  often  a  box  roll  arid  after 
that  diamond  and  oval  alternately. 

American  Wirebar  Sizes.— A  refinery  carries  about  six 
sets  of  standard  molds  and  varies  the  weight  of  the  bars 
cast  from  each  set  by  suitably  changing  the  depth  to  which 
the  molds  are  filled  within  the  limits  imposed  by  the  nec- 
essity of  keeping  a  bar  reasonably  square  in  cross-section. 
An  example  from  practice  is  shown  in  Table  42. 


TABLE  42. — TYPICAL  AMERICA*  WIREBAR  SIZES 


Weight, 
Ib. 

Length, 
in. 

Depth, 
in. 

Top  width, 
in. 

Bottom 
width,  in. 

135 

35K 

3% 

3M 

3^6 

175 

47K 

3% 

3% 

3K 

200 

49K 

3% 

4 

3^ 

225 

50 

4Ke 

4 

m 

250 

56 

3% 

m 

3% 

275 

55% 

4Ke 

4K 

*H 

300 

55K 

4% 

$4 

4^ 

REQUIREMENTS  OF  REFINED  COPPER          157 

Bars  of  135  Ib.  and  under  are  commonly  used  in  Europe, 
but  practically  not  at  all  in  the  United  States,  where  the 
usual  demand  is  between  175  and  225  Ib.  Very  heavy 
bars,  running  around  800  Ib.,  are  in  moderate  demand  on 
the  Continent  for  making  great  lengths  of  trolley  wire 
without  reverting  to  the  American  practice  of  brazing. 

Inasmuch  as  the  great  demand  for  power  in  a  rolling 
mill  is  in  the  first  two  or  three  passes,  the  roller  would  prefer 
a  long  bar  of  small  cross-section.  This,  however,  would 
increase  the  difficulties  in  casting  and  in  obtaining  sound 
bars,  and  developments  in  this  direction  are  improbable 
unless  it  be  in  connection  with  castings  made  on  end. 

Cakes  are  the  most  troublesome  class  of  castings  to 
make.  They  have  a  large  set  surface  and  this  is  rolled 
out  to  a  mirror-like  finish  where  every  flaw  is  readily  seen. 
Often  they  are  cast  on  " mixed"  rings,  so  that  wirebars  are 
interspersed  on  the  mold  ring  and,  as  perfect  pitch  for  a 
wirebar  is  not  necessarily  perfect  for  a  cake,  a  slight  dis- 
advantage results.  It  is  quite  impossible  to  cast  a  small 
ingot  and  a  heavy  cake  from  the  same  ladle  without  adjust- 
ment in  pitch,  as  the  shrinkage  and  gas  release  conditions 
are  so  different  in  the  two  cases. 

Occasionally  copper  is  shipped  as  cathodes  in  place  of 
ingots.  This  is  a  perfectly  logical  procedure,  as  the  fur- 
nace operation  is  not  a  true  refining  and  it  has  been  abund- 
antly shown  that  a  cathode  really  freed  from  electrolyte 
can  be  melted  to  high-grade  copper  without  any  furnace 
refining  whatever.  On  the  other  hand,  the  cathodes  must 
be  sheared  to  a  suitable  size,  they  are  often  brittle  and  there 
is  always  more  or  less  metal  loss  in  transit.  Aside  from 
nodules  being  dropped  by  the  wayside,  pieces  are  easily 
broken  off  by  thieves.  Also  cathodes  are  quite  variable 
in  individual  quality,  while  in  the  refining  furnace  thou- 
sands are  averaged  together.  The  result  is  that  ingots 
still  hold  their  own  in  the  brass  trade,  although  a  fair 
proportion  of  cathodes  are  sold  on  an  allowance  below  the 
price  of  wirebar  copper. 


CHAPTER  X 
COPPER  FROM  SECONDARY  MATERIAL 

In  addition  to  the  cirulating  products  incident  to  normal 
copper  refining,  the  plant  is  often  called  upon  to  treat  a 
wide  range  of  metallurgical'by-products  from  outside  sources 
as  well  as  miscellaneous  copper-bearing  scrap  material. 
Now  that  matte  smelting  plants  exist  on  the  Atlantic  sea- 
board, there  is  little  reason  to  consider  anything  in  the 
nature  of  ore  or  matte,  but  the  following  list  of  materials  is 
fairly  representative  of  what  offers: 

A.  Junk. 

a.  Lamps,  clocks,  etc. 

b.  Borings  and  chips. 

c.  Wire. 

d.  Sweepings. 

e.  Coins  and  alloys. 

B.  Mill  Products. 
a.  Wire. 

6.  Scale. 

c.  Cement. 

d.  Stampings. 

C.  Metallurgical  Products. 

a.  Cement. 

b.  Amalgamation  plates. 

c.  .Secondary  pig. 

d.  Nickeliferous  pig,  black  copper,  etc. 

e.  Furnace  bottoms. 

D.  Refinery  By-products. 

a.  Silver  refinery  slags. 

b.  Liberator  tank  cathode  products. 

c.  Flue  products. 

There  is  a  surprisingly  large  tonnage  of  such  miscellaneous 
material  available,  and  of  course  a  good  deal  of  it  is  ab- 
sorbed by  relatively  small  sweep  smelters,  who  generally 
feed  it  indiscriminately  into  a  small  blast-furnace,  making  a 

158 


COPPER  FROM  SECONDARY  MATERIAL         159 

foul  pig  which  is  sometimes  given  an  additional  reverbera- 
tory  treatment  and  sent  back  to  the  market  as  casting 
copper  and  sometimes  is  sold  to  an  electrolytic  refinery. 
Such  work  is  a  very  disturbing  element  in  a  copper  refinery 
unless  the  tonnage  accepted  is  great  enough  to  warrant  the 
establishment  of  regular  facilities  for  handling  it.  The 
general  questions  involved  are,  first,  sampling;  second  dis- 
position; third,  proper  furnace  construction;  fourth,  pos- 
sible saving  of  impurities  as  by-products. 

Sampling. — The  sampling  of  junk  is  an  art  in  itself.  The 
ordinary  junkman,  who  travels  around  hi  a  wagon  buying 
what  offers,  learns  in  time  to  make  very  close  estimates  of 
the  metal  values  of  such  material,  but  the  refinery  cannot 
afford  to  let  its  sampling  department  adopt  such  methods 
on  general  principles. 

The  first  rule  in  sampling  junk  is  to  classify  the  material. 
The  insulated  wire  is  put  in  one  lot,  the  bare  wire  in  another, 
the  alarm  clocks  and  automobile  lamps  in  a  third,  the 
sweepings  in  a  fourth,  the  borings  in  a  fifth,  and  so  on. 
Clocks  and  lamps  have  been  considered  as  a  class  because 
they  are  one  of  the  staples  of  the  junk  business.  Occasion- 
ally large  lots  of  Oriental  bronze  coins  come  into  the  market. 
A  noteworthy  instance  of  this  was  when  Japan  took  over 
the  control  of  Korea  and  exported  about  6,000,000  Ib.  of 
Korean  "cash." 

The  second  step  in  the  sampling  is  to  endeavor  to  get  a 
representative  part  of  each  classified  pile.  Corns  or  sweep- 
ings can  be  quartered  down  like  so  much  crushed  ore.  Wire 
can  be  clipped  here  and  there.  Borings  can  be  reasonably 
sampled  by  selection.  The  remainder  can  be  handled  only 
by  intelligently  taking  grab  samples,  and  this  requires 
experience  and  skill. 

The  third  rule  is  always  to  take  duplicate  samples  and  to 
hold  the  lot  intact  until  assays  are  out,  giving  an  oppor- 
tunity for  resampling  if  the  assays  show  sufficient  lack 
of  agreement  to  indicate  poor  sampling. 

The  next  step  varies  according  to  the  nature  of  the 
material.  With  dirty  metallic  material,  such  as  wire  or 


160  COPPER  REFINING 

borings  or  sweepings,  a  burning  loss  should  be  obtained  by 
soaking  with  gasoline  where  necessary  and  igniting,  noting 
the  weight  before  and  after.  With  miscellaneous  manu- 
factured material,  where  different  metals  are  used,  as  in  a 
clock,  or  containing  solder,  as  in  a  teapot,  the  only  way  is  to 
make  a  crucile  melt  of  a  rather  large  sample,  accounting  for 
the  bar,  slag  and  crucible  absorption.  Often  a  material 
has  to  be  sorted  into  metallics  and  dust,  as  in  the  case  of 
some  sweepings.  In  general,  therefore,  it  will  be  seen  that 
sampling  junk  is  a  tedious  and  expensive  business.  Most 
junk  carries  some  gold  and  silver  values  arising  from  jeweler's 
sweeps,  etc. 

Mill  products,  such  as  scrap  wire,  punchings,  etc., 
shipped  directly  from  the  wire  mill  or  other  factory  to  the 
refinery,  are  much  easier  to  sample.  Here  the  material  is 
already  classified  and  relatively  clean.  Care  must  be  taken 
to  see  that  no  bimetallic  wire  has  inadvertently  found  its 
way  into  straight  copper  wire  bundles.  It  would  be 
thought  that  blight  copper  wire  from  primary  sources 
would  need  no  sampling,  but  there  is  always  more  or  less 
grease  attached  to  it,  and  the  true  copper  content  is  nearly 
always  several  tenths  of  1  per  cent  below  standard.  Tinned 
wire  will  run  1  per  cent  or  more  low  and  bimetallic  wire 
may  run  40  per  cent  low.  The  refinery  in  turn  makes  a 
small  profit  in  that  the  wire  bars  sent  out  and  accepted  by 
the  trade  as  100  per  cent  copper  in  reality  run  seven  or 
eight  hundredths  of  1  per  cent  below  this. 

Among  the  metallurgical  products  two  require  special 
mention,  amalgamation  plates  and  furnace  bottoms.  Amal- 
gamation plates  are  copper  sheets  which  have  been  amalga- 
mated on  the  surface  and  used  to  collect  gold  and  silver 
values  from  finely  crushed  native  ores.  The  amalgam  has 
been  thoroughly  scraped  off  before  the  plate  is  scrapped  on 
account  of  wear  or  other  defect,  but  nevertheless  sufficient 
gold  is  retained  to  make  such  material  relatively  valuable. 
In  this  case  it  is  necessary  to  shear  the  plates  up  into  equal 
squares  of  a  few  inches  on  a  side;  then  take  systematically 
say  every  tenth  square,  following  such  a  system  of  reduc- 


COPPER  FROM  SECONDARY  MATERIAL         161 

tion  until  the  sample  is  small  enough  to  permit  of  crucible 
reduction. 

Furnace  bottoms  are  very  troublesome  to  handle.  They 
are  generally  received  in  large  masses  which  cannot  be 
drilled  or  broken.  They  are  usually  placed  in  a  furnace 
for  treatment  by  an  overhead  crane  at  a  time  when  the  roof 
is  out  for  rebuilding.  They  cannot  be  blasted  or  cut  apart 
and  the  only  way  to  reduce  the  size  of  a  large  piece  is  to 
build  a  small  flue  across  the  top  and  channel  with  a  hot 
oil  flame. 

The  only  way  to  sample  them  is  therefore  to  chip  around 
the  edges.  Unfortunately  they  are  often  quite  rich,  as 
silver  and  gold  tend  to  concentrate  in  furnace  bottoms. 
The  refiner  has  to  make  an  allowance  in  the  treatment 
terms  to  cover  the  sampling  danger  when  handling  such 
material. 

Disposition. — There  are  several  places  in  the  process 
where  secondary  material  can  be  introduced.  The  first 
is  the  slag  cupola,  the  function  of  which,  is  primarily  to 
treat  the  anode  and  wire  bar  furnace  slags  and  which  always 
has  capacity  to  spare.  Nearly  anything  in  the  scrap  line 
can  be  introduced  here,  but  it  does  little  more  than  extract 
the  metals  contained  and  put  them  into  compact  shape. 
Their  chemical  treatment  requires  oxidation,  which  can- 
not be  obtained  in  the  strongly  reducing  atmosphere  of  a 
black  copper  cupola.  Plants  treating  junk  exclusively 
usually  make  a  blast-furnace  melting  their  first  operation, 
using  some  briquetting  or  sintering  apparatus  to  take 
care  of  the  very  fine  material  before  charging. 

The  usual  point  of  entry  is  the  anode  furnace.  This 
insures  saving  the  silver  and  gold  contents  and  getting 
the  copper  in  electrolytic  form,  neither  of  which  is  possible 
to  the  small  junk  smelter. 

There  are  certain  objections  to  introducing  such  material 
into  an  ordinary  anode  charge.  In  the  first  place,  it  is 
usually  in  such  shape  physically  that  it  will  take  up  a  dis- 
proportionate amount  of  space  in  the  cold  charge.  This 
can  be  met  to  a  certain  extent  by  charging  it  at  the  re- 


162  COPPER  REFINING 

charge  period  when  there  is  plenty  of  room  in  the  furnace. 
Then  impure  material  delays  the  blowing  and  skimming, 
and  it  is  important  that  the  furnaces  operate  on  schedule 
time.  Finally,  quicker  work  and  better  elimination  of 
impurities  as  well  as  a  better  slag  concentration  for  possible 
by-product  recovery  are  made  by  treating  impure  material 
without  dilution.  This  leads  to  setting  an  anode  furnace 
apart  for  special  work  at  least  part  of  the  time. 
This  has  another  great  advantage  in  that  it  allows  direct 
costs  of  the  work  to  be  kept. 

Then  we  have  the  refining  or  wire  bar  furnaces.  Nothing 
should  be  added  to  a  cathode  charge  regarding  which 
there  can  be  any  question  as  to  quality.  It  is  customary 
to  include  as  admissible,  however,  scrap  wire  baled  in 
" cabbages"  which  originates  at  standard  wire  mills.  In 
ingot  charges,  where  conductivity  requirements  are  not 
quite  so  strict,  liberator  cathodes  from  the  first  row  of 
tanks  in  the  purifying  system  may  be  added  with 
discretion. 

Finally,  we  have  the  possibility  of  using  a  special  re- 
fining furnace  for  the  production  of  casting  copper.  This 
is  attractive  in  that  it  enables  material  which  is  practically 
free  of  silver  and  gold  and  high  in  copper — scrap  brass  for 
example — to  be  worked  directly  into  a  marketable  product. 
The  disadvantage  is  that  casting  copper  has  no  stable 
market.  It  always  sells  at  a  discount  and  at  times  is 
nearly  unsalable. 

Furnace  Construction. — Until  recent  years  the  standard 
refining  reverberatory  has  been  constructed  of  silica 
throughout.  In  some  special  cases  a  silica  brick  bottom 
has  been  employed,  but  the  usual  construction  was  silica 
sand  bottom  with  silica  brick  sides  and  roof.  Such  a 
furnace  is  built  with  a  cooling  vault,  cast-iron  bottom 
plates  mounted  on  brick  piers,  and  a  false  bottom  of  fire- 
clay brick.  The  vault  should  be  arranged  with  considera- 
tion of  a  possible  runout  of  copper  through  the  bottom. 
It  is  desirable  to  lead  the  molten  copper  away  from  the  vault 
before  it  sets,  so  that  it  can  be  recovered  without  tearing 


COPPER  FROM  SECONDARY  MATERIAL         163 

down  the  entire  furnace,  but  it  is  equally  essential  to  see 
that  it  does  not  run  into  the  casting  wheel  pit  or  ash  pit 
where  it  will  probably  come  into  contact  with  water  and 
cause  serious  explosions.  When  a  break-out  occurs  the 
proper  thing  to  do  is  to  charge  cold  copper  into  the  furnace 
to  set  the  charge,  which  is  not  difficult  on  account  of  the 
high  heat  conductivity  of  copper;  but  if  explosions  are 
occurring  it  is  not  safe  to  approach  the  furnace. 

A  sand  bottom  must  be  carefully  impregnated  with 
copper,  and  an  excellent  way  in  which  to  accomplish  this 
is  to  mix  the  sand  with  about  30  per  cent  of  its  weight  of 
copper  oxide  scale.  This  is  thoroughly  turned  over  and 
shoveled  into  the  previously  dried  furnace,  where  it  is 
calcined  at  a  moderate  heat  for  about  12  hr.  The  fire 
is  then  dropped  and  the  floor  carefully  leveled  and  tamped 
down.  The  furnace  is  then  fired  at  maximum  tempera- 
ture— perhaps  2700°F. — for  24  hr.  and  the  furnace  after 
cooling  is  ready  for  a  small  seasoning  charge  of  scrap 
wire  or  similar  light  material.  This  charge,  amounting  to 
perhaps  10  per  cent  of  the  rated  capacity  of  the  furnace, 
is  melted  and  ladled  out,  and  after  another  brief  cooling, 
one  twice  as  large  is  put  in.  After  this  has  been  ladled  the 
furnace  should  be  ready  for  service,  although  it  should  not 
be  crowded  for  a  few  charges,  starting  at  50  per  cent 
capacity  and  working  up. 

This  furnace,  even  when  treating  high-grade  material 
will  be  more  or  less  fluxed  by  the  cuprous  oxide  of  the 
charge,  the  silica  of  the  slag  being  taken  chiefly  from  the 
bottom  of  the  walls  near  the  metal  line.  The  old  treat- 
ment for  this  was  "fettling,"  or  plastering  corroded  spots 
with  fire  clay,  after  each  charge  was  cast. 

Prolonged  heating  in  the  treatment  of  foul  charges  is 
apt  to  smelt  some  of  the  bottom  of  the  furnace,  and  when 
pieces  of  the  bottom  begin  to  " float  up"  there  is  always 
danger  of  a  run-out.  Also  if  elements  such  as  lead  are 
present  in  appreciable  quantity  the  litharge  formed  will 
hungrily  attack  the  silica  of  the  furnace.  In  fact,  a  good 
silica  furnace  can  be  easily  ruined  by  a  single  charge  of 


164  COPPER  REFINING 

leady  copper.     More  than  2  per  cent  lead  is  always  cause 
for  great  anxiety. 

After  a  while  magnesite  and  chrome  brick  were  tried  as  a 
lining  material  for  the  walls  of  the  furnace  near  the  metal 
line.  Approved  practice  was  the  use  of  magnesite  for  the 
zone  protected  and  the  introduction  of  one  course  of  chrome 
brick  between  the  magnesite  and  silica  of  which  the  wall 
was  continued  up,  the  chrome  acting  as  neutral  material 
between  the  basic  and  acid  bricks. 

Magnesite  brick  spalls  badly  when  exposed  to  sudden 
temperature  changes.  Chrome  brick  is  not  subject  to  this 
objection,  but  fails  because  of  softening  when  exposed  to 
heavy  thrust  when  highly  heated.  Chrome  brick  is 
more  difficult  to  smelt  for  recovery  of  metal  values.1 

The  next  step  was  to  use  a  chromite  verb  with  great 
gain  in  heat  control  of  the  furnace,  since  small  dimensional 
changes  from  the  cutting  away  of  the  verb  cause  great 
differences  in  the  characteristics  of  a  furnace. 

Then  the  complete  basic  furnace  was  developed  with 
magnesite  bottom  and  chrome  walls  and  roof.2  The  sand 
bottom  furnace  is  constructed  with  the  bottom  inside  the 
walls;  the  basic  brick  furnace  is  built  with  the  walls  standing 
on  the  bottom  in  order  to  carry  the  thrust  of  the  inverted 
arch  directly  through  to  skewbacks  reacting  against  the 
brickstaves. 

With  complete  basic  furnaces  it  has  been  found  possible 
to  treat  material  high  in  lead,  and  which  may  require 
heavy  firing  for  several  days  without  damage  to  the  furnace, 
so  that  clean  anodes  may  be  made  from  foul  sources  of 
supply. 

By-Products. — Tables  43  and  44  give  an  idea  of  the 
character  of  some  of  the  low-grade  material  offering  in 
quantity. 

An  examination  of  these  analyses  shows  that  aside  from 
copper,  the  main  value,  there  are  often  present  considerable 
quantities  of  gold,  silver,  arsenic,  nickel,  cobalt,  lead,  zinc  and 

1  Pyne,  Trans.  A.  I.  M.  E.,  vol.  59,  p.  151. 

2  Addicks  and  Brower,  U.  S.  Patents  1083719  and  1148814. 


COPPER  FROM  SECONDARY  MATERIAL 

TABLE  43. — ANALYSES  OF  CEMENT  COPPER 


165 


Source 

Oz. 
gold 

Oz. 
silver 

Per  cent 
copper 

Per  cent 
iron 

Per  cent 
insoluble 

United  States  

Trace 

3.6 

80.93 

1.40 

8.80 

United  States  

0.01 

5.3 

71.53 

1.30 

17.06 

United  States  

0.28 

11.0 

60.58 

10.00 

1.68 

Australia 

0  45 

0  32 

78  28 

3  97 

3  50 

England 

0  05 

19  8 

76  62 

5  20 

1  63 

Belgium  

Trace 

0.40 

69.51 

8.50 

2.24 

Unknown 

Trace 

0  20 

32.18 

26  02 

5.99 

tin.  Where  gold  and  silver  are  present  in  sufficient  quantity 
to  pay  the  cost  of  electrolytic  treatment,  the  furnace  product 
should  obviously  be  anodes.  Where  the  silver  and  gold 
values  are  low,  casting  copper  may  be  considered  as  a 
product,  but  the  tune  required  for  refining,  market  deduc- 
tion for  casting  copper,  values  and  costs  must  be  carefully 
balanced  in  making  a  decision. 

Arsenic  can  be  almost  totally  removed  from  an  other- 
wise pure  bullion  as  sodium  arseniate  by  fluxing  with  soda 
ash.  This  method  is  employed  in  the  Lake  district,  the 
soluble  salt  being  leached  out  of  the  slag  with  water. 

When  any  of  the  other  impurities  are  present  in  quantity, 
we  may  consider  the  possibility  of  their  recovery  in  slags, 
in  flue  products  or  through  the  electrolyte  purifying  plant 
In  the  formation  of  slags  the  elements  oxidize  in  the  order 
of  their  basicity.  Then*  relative  affinities  for  oxygen  may 
be  judged  by  the  heats  of  formation  of  their  common  oxides, 
as  follows:1 


Values  from  Richards'  "Metallurgical  Calculations." 


Au2O3 -11,500 

Ag2O +  7,000 

CuO +37,700 

PbO +50,800 

NiO..                                   .  +61.500 


CoO +  64,100 

FeO +  65,700 

ZnO +  84,800 

SnO2 +141,300 


166 


COPPER  REFINING 


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COPPER  FROM  SECONDARY  MATERIAL         167 


From  these  figures  we  should  expect  to  get  first  the  tin, 
then  the  zinc,  then  the  group  of  cobalt-nickel-iron,  and 
finally  the  lead,  leaving  copper-silver-gold  in  the  metal 
bath.  In  practice,  however,  we  have  to  consider  the  fusi- 
bility of  these  oxides  and  the  readiness  with  which  they  form 
silicates.  The  situation  is  quite  different  from  ore  smelting 
in  that  we  are  dealing  with  metallic  alloys  and  always  in  the 
presence  of  a  preponderating  amount  of  copper. 

The  possibility  of  recovering  tin  -from  secondary  copper 
such  as  given  in  Table  44  is  indicated  in  Table  45  which 
gives  a  metal  balance  on  one  anode  furnace  charge. 

TABLE  45. — TIN  RECOVERY 
Input 


Material 

Weight 

Per  cent 
Sn 

Contents 

Crude  bullion       

53,567 

.     3.19 

1,709 

Crude  bullion 

50,874 

3  30 

1,679 

Crude  bullion 

61  125 

2  07 

1  265 

Crude  bullion                   .    .        .    . 

51,918 

2.23 

1,158 

Crude  bullion 

54,027 

2  10 

1,135 

Crude  bullion  

54,440 

3.02 

1,644 

Furnace  returns       .    .        

6,000 

0.27 

16 

Anode  scrap 

30,000 

0  15 

45 

Total                 

361,951 

2  39 

8,651 

Output 


Material 

Weight 

Per  cent 
Sn 

Contents 

Per  cent 
Sn  input 

Anodes 

278  801 

0  61 

1  701 

19  7 

First  slag  

7,686 

0  38 

29 

0  3 

Second  slag  
Third  slag 

17,796 
39,657 

0.78 
14  43 

139 
5,722 

1.6 
66  2 

Metallics  
Unaccounted  for  

4,019 
13,992 

0.61 

25 
1,035 

0.3 
11.5 

Total 

361  951 

8,651 

ioo  o 

168 


COPPER  REFINING 


There  has  been  collected  in  the  third  slag  66  per  cent  of 
the  tin.  If  we  take  the  copper  in  this  slag  at  25  per  cent,  we 
shall  have  upon  reduction  to  metal  a  foul  alloy  running  30 
to  35  per  cent  in  tin,  a  reasonable  first  step  toward  the 
recovery  of  this  metal. 

Zinc  is  not  a  sufficiently  valuable  metal  to  warrant  the 
cost  of  its  extraction  from  impure  or  low-grade  products. 
Inasmuch  as  a  copper  furnace  is  operated  at  a  temperature 
above  the  boiling  pint  of  zinc,  it  would  be  possible  to  make  a 
certain  recovery  as  fume,  but  as  this  would  be  contaminated 
with  other  volatile  impurities  such  as  arsenic  and  antimony 
present  in  foul  bullion,  such  an  operation  would  have  no 
commercial  standing. 

While  iron,  cobalt  and  nickel  stand  close  together  in  the 
heat  of  oxidation  scale  they  show  a  marked  difference  in 
slagability.  In  fact  the  ease  of  removal  in  a  reverberatory 
furnace  might  be  arbitrarily  indicated  as  1.00  for  iron,  0.10 
for  cobalt,  and  0.01  for  nickel. 

This  is  illustrated  in  a  general  way  by  some  published1 
data  of  experiments  conducted  in  a  2-ton  furnace  at  the 
Wysk  smelter  in  Russia,  some  of  the  results  of  which  are 
shown  in  Table  46. 

Iron  is  of  course  far  too  cheap  a  metal  to  be  considered 
as  a  by-product  except  in  very  large-scale  ore  smelting 

TABLE  46. — RELATIVE  SLAGABILITY  OF  IRON,  COBALT  AND  NICKEL 


After 

After 

Original 

first 

second 

bullion 

refining 

refining 

Copper,  per  cent  

99.55 

99.42 

99.66 

Iron,  per  cent 

3  04 

0  Oil 

0  008 

Cobalt,  per  cent  

0.894 

0.023 

0.024 

Nickel,  per  cent  

0.408 

0.110 

0.109 

Iron.  . 

1  000 

0  004 

0  003 

Cobalt  

1.000 

0.026 

0.027 

Nickel  

1.000 

0.270 

0.267 

Hof man's  Metallurgy  of  Copper,  p.  391. 


COPPER  FROM  SECONDARY  MATERIAL         169 


operations.  It  can  be  easily  removed  from  copper  bullion 
by  slagging,  but  not  without  the  coincident  removal  of  a 
small  amount  of  the  copper  as  cuprous  oxide,  which  forms 
the  scorifying  agent. 

Cobalt  can  be  removed  with  reasonable  facility  in  like 
manner  and  collected  in  a  slag  running  perhaps  20  per  cent 
cobalt  and  20  per  cent  copper,  but  truly  cobalt-free  copper 
could  not  be  produced  without  slagging  a  large  amount  of  the 
copper  in  repeated  scorifications. 

Nickel  is  much  more  difficult  to  deal  with,  and  is  generally 
recovered  by  separation  as  a  soluble  sulphate  in  electrolytic 
refining.  Considerable  data  are  available  on  its  behavior  in 
the  anode  furnace.  An  example  of  some  large-scale  ex- 
periments is  given  in  diagrammatic  form  in  Fig.  33,  and 
these  curves  are  analyzed  in  Table  47. 


0     10    20   30   40    50    60  70    80  90    0      1     2     3     4      56     7     8     9     10 
Per    Cent. 

FIG.  33. — Behavior  of  nickel  in  the  copper  reverberatory. 
TABLE  47. — BEHAVIOR  OF  NICKEL  IN  THE  COPPER  REVERBERATORY 


Per  cen 

t 

Nickel  in  charge  

0  0 

0  5 

1  0 

1.5 

2  0 

Elimination  

37  2 

57  0 

70  0 

80  4 

Nickel  in  product 

0  0 

0  32 

0  44 

0  45 

0  40 

Slag 

2  0 

3  8 

5  7 

7  6 

9  4 

Nickel  in  slag  

0  0 

4  9 

10  0 

13  8 

17  1 

170  COPPER  REFINING 

^ 

It  is  evident  that  more  prolonged  working  of  the  charges 
richer  in  nickel  would  result  in  a  somewhat  lower  nickel  in 
the  product  and  a  larger  quantity  of  slag  lower  in  grade 
as  to  nickel,  but  is  it  quite  out  of  the  question  to  reduce 
the  nickel  tenor  of  a  large  charge  to  the  degree  shown  by 
the  small  Wysk  furnace.  The  slag,  if  treated  in  a  matting 
furnace  for  the  recovery  of  copper,  will  lose  a  large  part  of 
its  nickel  to  the  new  slag;  if  treated  in  a  black  copper  cupola, 
the  nickel  will  be  nearly  all  reduced  with  the  copper. 

Lead  is  readily  separated  from  molten  copper  as  litharge. 
As  this  active  base  rapidly  attacks  siliceous  material,  quite 
small  quantities  of  lead  demand  a  basic  furnace  for  safe 
handling.  Lead  is  practically  insoluble  in  pure  copper, 
but  in  the  presence  of  sulphur  or  various  impurities  it  is 
absorbed,  and  considerable  quantities  are  found  in  secon- 
dary copper  and  occasionally  in  blister  where  leady  copper 
ores  have  been  treated.  Certain  bearing  metals  are  made 
of  a  mixture  chiefly  of  copper  and  lead,  and  the  black 
Korean  coins  listed  in  Table  44  showed  high  percentages  of 
lead.  When  melting  down  large  quantities  of  these  in  a 
furnace  with  a  basic  brick  bottom  placed  over  a  cooling 
vault,  it  was  noticed  that  practically  pure  lead  sweated 
through  the  bricks  underneath,  indicating  liquation  and  a 
possible  means  of  separating  lead  from  copper.  A  copper- 
lead  slag  which  can  be  profitably  smelted  for  both  metals  is 
readily  made  from  such  material. 

In  general,  electrolytic  refineries  recover  a  large  part  of 
any  nickel  or  lead  entering  their  plants  in  sufficient  quanti- 
ties, and  steps  would  be  taken  to  recover  cobalt  and  tin 
did  the  crude  copper  carry  sufficient  of  these  metals,  but 
the  disturbance  of  the  process  due  to  interference  of  these 
impurities  in  the  main  business  of  copper  refining  over- 
shadows the  possibility  of  placing  impurities  on  a  revenue- 
producing  basis  except  where  large  regular  quantities  have 
to  be  dealt  with.  Where  casting  copper  is  made,  a  por- 
tion of  these  alloy  impurities  remains  and  actually  improves 
the  copper  for  certain  purposes  by  increasing  its  fluidity 
or  altering  its  color. 


CHAPTER  XI 
THE  POWER  PROBLEM 

An  electrolytic  copper  refinery  may  obtain  its  power 
from  a  self-contained  steam  or  gas  plant,  from  an  adjacent 
water  power  or  by  transmission,  but  while  each  case  will 
call  for  a  different  type  of  power  plant  as  far  as  the  provi- 
sion of  current  for  the  tanks  is  concerned,  the  demand  for 
fuel  for  smelting  and  for  steam  for  heating  liquors  will 
remain  unchanged.  The  fact  that  both  fuel  and  steam  are 
required  introduces  waste  heat  boilers  in  any  event  and 
most  refineries  have  steam  driven  electrolytic  generators, 
although  there  are  notable  exceptions.  The  conditions 
surrounding  each  case  call  for  individual  study  of  all  the 
factors  entering  and  it  is  the  purpose  here  merely  to  point 
out  the  general  requirements  and  limitations  of  the 
problem  as  a  whole. 

In  the  early  plants,  before  the  days  of  waste-heat  boilers 
and  with  live  steam  heating,  the  total  water  entering  the 
plant  could  be  considered  as  utilized  roughly  one-third 
for  generation  of  electrolytic  power,  one-third  for  steam  for 
heating  and  miscellaneous  uses  and  one  third  for  make-up 
water  in  the  tanks  and  at  the  boshes.  The  ideal  steam 
driven  plant  would,  therefore,  be  one  where  the  steam  gene- 
rated by  waste  heat  at  the  reverberatories  would  be  suffi- 
cient to  supply  the  demands  of  the  engine  room,  the 
exhaust  from  the  engine  room  was  used  for  heating  and  the 
condensate  finally  was  used  as  make-up  water  at  either  the 
boshes  or  the  boilers.  This  ideal  solution  is  yet  far  from 
being  reached  but  the  very  uniform  conditions  of  load  and 
output  under  which  a  refinery  is  run  has  enabled  sufficient 
progress  to  be  made  in  this  direction  considerably  to 
handicap  hydroelectric  and  gas  plant  competition. 

171 


172 


COPPER  REFINING 


Waste  Heat  Boilers. — It  is  perhaps  most  logical  to  start 
with  the  consideration  of  the  amount  of  steam  available 
as  a  by-product  at  the  reverberatories. 

Given  a  fair  grade  of  bituminous  coal  a  modern  refining 
furnace  will  average  eight  to  nine  tons  of  product  per  ton 
of  coal  burned.  As  there  are  two  furnace  operations  and 
as  the  anode  furnace  treats  some  15  per  cent  of  anode 
scrap  and  various  other  items  contribute  to  swell  the  net 


•  XI 


0 


8 


16 


20 


12 
Hours 

FIG.  34. — Steaming  rate  of  waste  heat  boilers. 


tonnage  treated  in  either  furnace,  a  ratio  of  four  to  one  is 
safe,  giving  25  per  cent  fuel  consumption  on  a  cathode  basis. 
A  properly  designed  boiler,  placed  close  to  the  throat  of 
the  furnace  and  equipped  with  an  economizer  can  be  de- 
pended upon  to  give  an  average  equivalent  evaporation  of 
between  six  and  seven  pounds  of  water  per  pound  of  coal 
fired.  We,  therefore,  have  available  about  1.5  Ib.  of  steam 
per  pound  of  cathode  output,  or  in  other  terms  a  200- 
ton  furnace  will  develop  at  least  363  boiler  horse-power. 


THE  POWER  PROBLEM  173 

The  rate  of  steaming  is  not  uniform  as  although  the  coal 
in  a  large  modern  furnace  is  consumed  quite  steadily  through- 
out the  twenty-four  hours,  the  two  extremes  of  poling  and 
charging  give  large  differences  in  the  temperature  of  the 
gases  leaving  the  furnace.  The  situation  is  graphically 
expressed  in  Fig.  34.  Where  a  number  of  furnaces  are 
feeding  boilers  in  parallel  their  operations  are  usually 
staggered  enough  to  give  a  sufficiently  uniform  steam  supply 
to  avoid  any  serious  irregularities  in  the  demand  upon  the 
main  boiler  installation.  Where  the  electrolytic  power  is 
obtained  by  transmission  this  variable  supply  might  cause 
undesirable  fluctuations,  but  in  this  case  most  of  the  steam 
would  be  used  for  heating  purposes  where  considerable  va- 
riations are  temporarily  permissible.  On  Sundays  the 
waste  heat  steam  supply  is  lessened  but  so  is  the  miscella- 
neous demand. 

A  good  general  discussion  of  the  whole  question  of  waste 
heat  boiler  equipment  has  been  given  in  a  paper  by  Arthur 
D.  Pratt,1  and  of  the  particular  application  to  copper 
refining  furnaces  in  an  article  by  Clarence  L.  Brower.2 
The  secret  of  successful  operation  lies  in  placing  the  boiler 
as  close  as  possible  to  the  furnace  throat  and  in  the  exclu- 
sion of  air  infiltration.  A  good  modern  installation  ob- 
tains a  temperature  of  1600°  to  1700°F.  in  the  boiler 
fire-box  as  against  1000°  to  1200°  in  the  earlier  attempts. 
The  exit  gases  may  be  brought  down  to  500°F.  or  lower. 

Electrolytic  Power. — As  has  been  discussed  in  Chapter 
IV,  the  power  required  in  the  tank  house  depends  upon  the 
current  density  and  the  temperature  of  the  electrolyte.  In 
average  practice  the  density  will  be  around  18  amperes 
per  square  foot  and  the  temperature  about  130°F., 
in  which  case  the  power  consumption  will  approximate 
350  kw.-hr.  per  ton  of  copper  deposited.  In  general  a 
refinery  with  a  capacity  of  100,000  tons  a  year  may  be 
considered  large  enough  to  be  a  thoroughly  economical 
unit  and  this  would  call  for  a  dependable  power  capacity 

1  Trans.' A.  S.  M.  E.,  vol.  38,  p.  599. 

2  E.  &  M.  J.,  vol.  99,  p.  892. 


174  COPPER  REFINING 

of  about  4000  kw.  Whatever  type  of  steam  driven  prime 
mover  be  adopted,  a  thoroughly  modern  power  plant  of 
this  size  should  deliver  a  kilowatt-hour  at  the  switchboard 
for  not  over  15  Ib.  of  steam,  given  reasonable  superheat 
and  vacuum  and  excluding  auxiliaries.  We,  therefore,  re- 
quire at  the  throttle  roughly  2.6  Ib.  of  steam  per  pound  of 
copper  refined,  or  nearly  double  the  1.5  Ib.  available  from 
the  waste  heat  boilers. 

Choice  of  Prime  Mover. — In  cases  where  water  power  is 
cheaper  than  steam  at  the  place  where  a  refinery  has  been 
located,  generally  for  reasons  quite  apart  from  the  cost  of 
power,  it  is  of  course  employed  and  we  have  such  a  case 
in  the  refinery  at  Great  Falls,  Montana.  In  the  same  way 
transmitted  power  originating  at  a  water  power  is  used  at 
Tacoma  and  to  a  certain  extent  at  Baltimore.  A  very 
unusual  instance  of  power  transmitted  from  a  steam  gen- 
erating plant  is  that  of  the  electrolytic  department  of  the 
leaching  plant  at  Chuquicamata,  Chile.  In  general  a 
refinery  is  located  with  reference  to  copper  freights  rather 
than  cost  of  power  and  in  most  locations  steam  power  is 
cheaper  than  water  power. 

Gas  power  has  also  been  more  or  less  discussed  but  up  to 
the  present  time  nowhere  applied  to  copper  refining  work 
as  far  as  power  generation  is  concerned.  The  great  first 
cost  of  the  installation  and  its  low  economy  at  partial  loads 
have  always  been  handicaps.  The  history  of  refineries 
has  been  that  they  are  continually  being  added  to  and  ex- 
panded and  that  even  individual  units  are  seldom  operated 
under  the  conditions  for  which  they  were  designed,  while 
above  all  else  a  gas  plant  lacks  elasticity.  On  a  basis  of 
thermal  efficiency  some  very  interesting  figures  may  be 
worked  up,  but  very  large  steam  turbine  units  can  now 
practically  equal  these. 

As  between  steam  turbines  and  reciprocating  engines, 
the  question  is  largely  a  question  of  size  of  unit. .  A  turbine 
can  expand  steam  down  to  the  highest  attainable  vacuum 
without  unreasonable  increase  in  size  and  owing  to  its 
high  rotative  speed  very  large  units  can  be  built  in  small 


THE  POWER  PROBLEM 


175 


compass.  In  the  case  of  a  reciprocating  engine  the  speed 
is  limited  by  valve  gear  considerations  while  cylinders 
larger  in  diameter  than  54  in.  are  undesirable  and  the 
practical  limit  of  size  is  about  72  in.  diameter  by  72  in. 
stroke.  The  low  pressure  cylinder  is  frequently  divided 
into  two  in  making  a  four  cylinder  triple  expansion  unit. 
These  considerations  bring  us  to  about  2000  horse-power 
discharging  into  27^  in.  vacuum  as  a  reasonable  limit 


20 
19 

18 

I 
ft 

816 

1 

01 

I'5 

1 

14 
13 

150°lbjGage  Pressure 
60  E.  Superheat 

27Vi"Vamiim 

\ 

^c 

/ 

S^ 

^C 
^^Cp 

Impound 

v**£ 

^ 

_TwSS 

£ylinde 

r—  Conn 

/ 

>iTlTd     w 

X 

•sfiss. 

Cylind 

„& 

v/ 

800          900          1000         1100  1200         1300         1400        1500 

Kilowatt  Output 
FIG.  35. — Water  rate  of  1200  KW.  unit. 

to  size  of  unit  for  a  reciprocating  engine  and  the  steam  econ- 
omy attainable  is  indicated  in  Fig.  35.  A  turbine  of  this 
size  can  hardly  equal  this  performance  even  with  higher 
vacuum  and  when  the  efficiencies  of  conversion  from  alter- 
nating to  direct  current  are  included  it  is  under  a  serious 
handicap  insofar  as  economy  of  steam  consumption  is 
concerned.  In  larger  sizes,  however,  the  turbine  reaches 
very  high  economy.  Figure  36  shows  approximately  what 
water  rate  may  be  expected  with  170  Ib.  gage  pressure, 
150°F.  superheat,  28^  in.  vacuum  after  allowing  7  per  cent 
loss  for  a.c.-d.c.  conversion.1 

1  Trans.  A.  I.  E.  E.,  vol.  xxxiii,  p.  1133. 


176 


COPPER  REFINING 


The  advantages  of  a  turbine  are  based  partly  upon  its 
compactness  and  its  ability  to  utilize  high  superheat  and 
vacuum.  One  inch  of  vacuum  or  50°F.  superheat  will 
vary  the  water  rate  five  or  six  per  cent  in  a  high  pres- 
sure turbine.  As  to  the  space  factor  it  must  be  remembered 
that  its  value  is  considerably  less  at  a  refinery  than  in  a 
city  power  house  and  regarding  condensation  the  cost  of 


24 


20 


16 


12 


170  Ib.Gage  Piessure 

150  F.  Superheat 

28  "  Vacuum 


former  Plus  Rotary 


246  8  10  ,12. 

Capacity  of  Unit  in  Thousands  of  Kw. 

FIG.  36. — Steam  turbine  water  rates. 


14 


producing  high  vacuum  must  not  be  lost  sight  of.  The 
temperature  of  saturated  steam  drops  very  rapidly  as 
vacuum  increases  and  if  cold  water  is  not  available  the  cost 
of  circulating  water  may  become  very  great.  These  con- 
ditions are  illustrated  by  the  curves  in  Fig.  37. 

The  whole  question  of  the  relative  value  of  the  several 
types  of  prime  mover  has  been  reviewed  by  Stott,  Pigott 
and  Gorsuch,1  and  the  special  application  to  electrolytic 
work  was  discussed  at  a  joint  meeting  of  three  of  the  engi- 

1  Trans.  A.  I.  E.  E.,  vol.  xxxiii,  p.  1133. 


THE  POWER  PROBLEM 


177 


neering  societies  several  years  ago.1  In  a  general  way  it 
can  be  stated  that  with  cheap  coal  the  plant  will  be  steam 
driven,  using  reciprocating  engines  if  small  and  turbo- 
alternators  distributing  through  rotary  converters  if  large. 
With  expensive  coal,  a  fairly  large  plant  and  regular  ton- 
nage a  gas  plant  could  be  made  to  pay.  With  water  power 
close  at  hand  it  would  probably  displace  either  steam  or  gas. 


10  F.  Loss  in  Tail  Pipe 

"  Vacuum  Loss 

in  Exhaust  Pipe 


50 


55  60  65  70  75 

Temperature  of  Water,  Deg  E. 
FIG.  37. — Water  required  for  condensing. 


The  Electric  Generator. — The  tank  house  circuits  de- 
mand (a)  direct  current,  (b)  high  amperage,  (c)  low  voltage 
and  (d)  wide  voltage  range.  The  amperage  used  has 
steadily  increased  as  the  generator  builders  have  found  it 
possible  to  build  larger  capacity  commutators  and  to-day 
runs  between  10,000  and  15,000  amp.  per  circuit. 

1  Newbury,  Trans.  A.  I.  E.  E.,  vol.  xxxiii,  p.  1;  Longwell,  Jour.  A.  S.  M.  E., 
July,  1914;  Addicks,  Trans.  A.  E.  S.,  vol.  xxv,  p.  65. 
12 


178  COPPER  REFINING 

The  voltage  is  generally  limited  to  200  volts  or  less  per 
circuit  and  the  maximum  unit  to  be  considered  is,  there- 
fore, 3000  kw.,  while  10,000  amp.  at  120  volts  or  1200 
kw.  would  more  nearly  represent  present  practice.  A 
1,200-kw.  generator,  direct  connected  to  a  reciprocating 
engine,  would  call  for  1,900  i.h.p.,  which  is  about  the 
limiting  engine  size. 

The  generator  problem  is  almost  wholly  one  of  commuta- 
tion. This  subject  has  been  reviewed  at  length  under  the 
heading  of  "  Physical  Limitations  in  D.C.  Commutating 
Machinery"  by  B.  G.  Lamme.1  The  simplest  case  is  that 
of  the  direct  connected  reciprocating  engine.  Here  the 
slow  engine  speed,  certainly  not  over  125  r.p.m.,  permits 
reasonable  freedom  in  commutator  design,  and  the  volts 
per  segment,  length  of  bar  and  amperes  per  brush  arm 
can  all  be  placed  at  conservative  figures.  It  must  be 
remembered  that  steady  duty  at  100  per  cent  current  load 
and  occasionally  at  abnormally  low  voltages  imposes  corn- 
mutating  conditions  far  more  severe  than  met  with  in 
average  practice  and  that  any  roughening  of  the  commuta- 
tor' surface  becomes  rapidly  worse.  Commutating  pole 
field  construction  has  made  a  great  advance  toward  spark- 
less  commutation  in  all  electrolytic  generators. 

The  efficiency  of  a  direct  connected  engine  type  generator 
exclusive  of  bearing  friction  will  be  about  94  per  cent  and 
the  e.h.p.  delivered  by  the  generator  should  be  about  88 
per  cent  of  the  i.h.p.  of  the  engine.  The  total  losses  on  a 
750-kw.  C-W.  generator  built  in  1905  are  given  in  Table 
48. 

The  direct  current  steam  turbine  has  also  given  satisfac- 
tion in  service,  although  there  is  a  practical  limit  to  the  size 
of  commutator  which  can  be  constructed,  as  it  is  necessary 
to  restrict  the  diameter  owing  to  the  high  rotative  speeds 
required.  A  two  cylinder  compound  driving  a  direct  con- 
nected generator  and  exhausting  into  a  direct  current  tur- 
bine connected  electrically  in  parallel  makes  a  satisfactory 
unit  and  several  such  pairs  are  in  successful  operation.  The 

1  Trans.  A.  I.  E.  E.,  vol  xxxiv,  p.  1739. 


THE  POWER  PROBLEM 

TABLE  48 


179 


Voltage,  volts 

125 

125 

100 

Current  amperes 

0 

6,000 

9,500 

Output,  kw.  .  .                              

0 

750 

950 

Bearing  friction,  watts  

4,000 

4,000 

4,000 

Windage,  watts  

1,600 

1,600 

1,600 

Brush  friction,  watts  
Core  and  pole  losses,  watts 

6,200 
18,000 

6,200 
18,500 

6,200 
15,000 

Field  copper,  watts  

6,560 

7,930 

4,680 

Armature  copper  watts 

11,750 

29,400 

Brush  resistance  losses,  watts 

8,650 

15,400 

Total  losses,  watts  

36,360 

58,630 

76,280 

Mechanical  losses,  per  cent  input  
Magnetic  losses,  per  cent  input  

32.4 
49.4 

1.5 
2.3 

1.1 
1.5 

Electrical  losses,  per  cent  input  
Output  (efficiency),  per  cent'  input  . 

18.2 
0.0 

3.5 
92.7 

5.0 

92  4 

Load,  per  cent  rating  

0.0 

100.0 

127.0 

steam  economies  possible  with  such  combinations  are 
indicated  in  Fig.  35.  This  same  field  is  covered  by  geared 
units,  a  rather  high-speed  generator  being  geared  to  a  rather 
low-speed  turbine.  For  large  plants  electrical  rather  than 
mechanical  connection  between  turbine  and  generator,  as 
provided  by  rotaries  has  many  advantages. 

The  unipolar  generator  which  delivers  direct  .current  with- 
out a  commutator  and  which  may  be  driven  at  turbine 
speeds  is  a  very  pretty  theoretical  solution  of  the  whole 
problem.  Practically,  while  commercial  units  have  been 
built1  many  special  problems  are  involved  in  the  collection 
of  the  current  and  this  type  of  apparatus  has  not  received 
any  considerable  application. 

If  a  turbo  alternator  is  used  as  the  main  source  of  current 
allowance  must  be  made  for  the  efficiency  of  the  distribut- 
ing system,  the  transformers,  conductors  and  rotary  con- 
verters absorbing  about  7  per  cent  of  the  A.C.  output. 

The  voltage  regulation  should  be  made  wide  and  is  satis- 
factorily accomplished  by  hand  control.  In  the  case  of 
reciprocating  engines  the  speed  may  be  altered  so  as  to  keep 

,  Transaction  A.  I.  E.  E.,  vol.  XXXI,  p.  1811. 


180  COPPER  REFINING 

the  engine  operating  at  the  best  cut-off.  This,  however, 
means  some  adjustable  attachment  to  the  engine  governor 
which  has  its  draw-backs.  With  ordinary  direct  current 
generators,  either  engine  or  motor  driven,  the  usual  hand 
rheostat  gives  satisfactory  voltage  control  and  modern 
commutating  pole  generators  will  operate  well  at  surpris- 
ingly low  voltages.  In  the  case  of  rotary  converters  it  is 
necessary  to  use  booster  control  on  the  alternating  current 
side,  which  may  be  supplemented  by  employing  multi-tap 
transformers. 

A  circuit  of  120  volts  would  probably  have  360  tanks 
connected  in  series  and  when  drawing  copper  two  groups  of 
say  36  tanks  each  might  be  cut  out.  This  would  call  for  a 
20  per  cent  cut  in  the  normal  operating  voltage  during  this 
period.  On  the  other  hand  while  a  tank  house  is  supposed 
to  be  operated  on  a  constant  current  basis,  this  is  very  apt 
to  become  constant  wattage,  the  current  density  being 
temporarily  increased  during  periods  of  low  voltage. 

Heating  Electrolyte.  Heating  the  electrolyte  has 
already  been  discussed  in  Chapter  IV.  At  what  tempera- 
ture it  will  pay  to  carry  the  liquor  must  be  determined  for 
each  case.  Eastern  refineries  expend  about  0.75  Ib.  of 
steam  per  pound  of  copper  produced  to  raise  this  temperature 
say  30°F.  above  the  temperature  which  would  result  from 
electrical  heating.  This  means  a  power  saving  at  the 
generator  of  15  per  cent,  taking  an  overall  temperature 
coefficient  of  0.5  per  cent  per  degree  Fahrenheit.  If  we 
are  using  3.0  Ib.  of  steam  (including  auxiliaries)  per  pound 
of  copper  at  the  engines  the  saving  will  be  0.45  Ib.  as  against 
0.75  Ib.  required  at  the  heating  coils  and  it  is  evident  that 
live  steam  heating  would  not  pay.  If  we  run  an  electrolytic 
unit  non-condensing  it  will  take  about  4.0  Ib.  of  steam 
instead  of  3.0.  If  we  are  able  to  run  a  proper  proportion 
of  the  circuits  non-condensing  we  can,  therefore,  furnish 
steam  for  heating  the  tanks  and  obtain  an  average  overall 
steam  rate  of  3.2  Ib.  on  85  per  cent  of  the  original  resistance 
or  2.7  Ib.  per  Ib.  of  copper  instead  of  3.0  when  running 
without  heating. 


THE  POWER  PROBLEM  181 

Where  the  electrolytic  power  is  hydraulic  live  steam  heat- 
ing may  not  pay  if  there  is  not  waste  heat  boiler  capacity 
in  excess  of  other  needs,  but  except  where  very  high  densi- 
ties are  employed,  thereby  predicating  clean  bullion  and 
resulting  in  considerable  electrical  heat  being  generated  in 
the  tanks,  it  generally  is  desirable  to  heat  the  electrolyte 
from  purely  metallurgical  considerations. 

The  condensate  from  heating  coils  is  generally  collected 
for  use  in  washing  the  anodes  and  cathodes  free  from  slimes, 
bluest  one  crystals,  etc.,  and  a  small  amount  of  heat  is  in 
this  way  returned  to  the  system  as  well  as  a  safeguard 
established  against  leaky  heating  coils  withdrawing  any 
electrolyte  and  sending  it  to  the  sewer. 

Boiling  Tanks. — The  silver  refinery  and  the  purifying  sys- 
tem or  bluestone  plant  will  always  require  a  certain  amount 
of  boiling  or  evaporating  This  varies  greatly  in  amount  in 
different  plants.  In  a  very  general  way  it  may  be  stated 
that  there  wrill  be  two  pounds  of  steam  condensed  for 
every  pound  of  water  evaporated  and  that  about  10  per 
cent  of  the  total  steam  used  will  be  employed  in  this  way 
in  the  average  plant. 

Steam  siphons  used  for  pumping  are  really  heaters,  their 
efficiency  as  pumps  running  but  a  few  hundredths  of  one  per 
cent.  The  steam  consumption  of  a  siphon  depends  upon 
the  original  design  and  upon  the  amount  of  nozzle  wear 
at  the  time  of  test,  but  one  per  cent  dilution  of  the  liquor 
pumped  is  probable  and  three  per  cent  possible. 

Heating  Feed  Water. — The  steady  full  load  conditions  of 
the  electrolytic  engines  call  for  a  complete  installation  of 
economizers.  Probably  ten  per  cent  of  the  steam  fur- 
nished by  the  boilers  will  be  called  for  as  exhaust  steam  for 
further  heating  of  the  feedwater. 

Miscellaneous  Demand. — There  is  a  considerable 
amount  of  pumping  to  be  done,  such  as  circulating  elec- 
trolyte, circulating  bosh  water,  supplying  condensers,  fire 
system,  general  water  supply,  feeding  boilers  and  high  pres- 
sure system  for  hydraulic  cylinders.  There  is  a  more  or 
less  extensive  compressed  air  system  for  operating  copper 


182 


COPPER  REFINING 


hoists  and  pneumatic  tools.  There  is  power  required  for 
the  motor  drives  of  cranes,  shop  tools,  ventilating  fans, 
sampling  drills  and  possibly  an  electric  railway.  Finally 
there  is  the  lighting  system. 

If  steam  driven,  non-condensing,  in  a  large  plant,  these 
various  units  will  consume  from  one-third  to  one-half  as 
much  steam  as  is  called  for  by  the  electrolytic  generators 
operated  condensing.  This  steam  may  advantageously  be 
turned  into  the  heating  systems,  the  requirements  of 
which  it  roughly  fills,  turning  all  of  the  steam  from  elec- 
trolytic units  into  the  condenser.  Another  solution  is  to 
electric  drive  all  auxiliaries,  supply  the  current  by  means  of 
a  non-condensing  light  and  power  unit.  There  is  even  the 
opportunity  of  utilizing  an  electrolytic  spare  for  this  ser- 
vice, thereby  reducing  the  total  equipment  called  for,  the 
plant  consisting  of  several  electrolytic  generators,  one 
light  and  power  unit  and  one  spare  for  both  services. 

Summary. — The  total  steam  consumption  of  a  copper 
refinery  will  vary  within  wide  limits,  depending  upon  the 
size  of  plant,  the  current  density  employed  and  the 
amount  of  purifying  of  the  electrolyte  required  and  may  be 
stated  somewhat  as  in  Table  49. 


TABLE  49 


Lb.  steam 
per  Ib. 
copper 


A.  Electrolytic  power,  condensing 

B.  Light,  power,  water,  air  and  auxiliaries,  non-condensing 

C.  Evaporating,  live  steam . 

D.  Heating,  exhaust  steam 

E.  Total  steam  demand,  A  +  B  +  C  +  D 

F.  Exhaust  steam  for  D  from  B 

G.  Contributed  by  waste  heat  boilers 

H.  Net  steam  demand,  E  -  (F  +  G] : 


2.5to3.0 
1.0  to  1.5 
0.5to  2.0 
l.Oto  1.5 
5.0to  8.0 
l.Oto  1.5 
1.5  to  1.5 
2.5to5.0 


CHAPTER  XII 
ELEMENTS  OF  DESIGN 

Up  to  the  present  time  electrolytic  refining  has  been  con- 
ducted almost  exclusively  in  the  United  States,  some  eighty 
per  cent  of  the  world's  production  being  there  treated. 
This  has  been  due  to  several  factors.  The  North  American 
continent  has  produced  the  great  bulk  of  the  world's  output 
of  copper,  metallurgical  operations  have  been  developed 
on  a  larger  scale  in  America  than  elsewhere  and  the  local 
copper  production  has  been  largely  consolidated  into  a  few 
large  groups  each  capable  of  supporting  one  or  more  large 
refineries. 

Location. — Since  the  war  the  nations  of  Europe  have 
developed  a  new  desire  to  control  key  industries  complete 
within  their  own  borders,  home  or  colonial,  and  new  tariffs, 
trade  routes,  etc.,  are  likely  to  cause  the  building  of  various 
new  refineries.  In  general  the  factors  associated  with  the 
movement  of  copper  from  mine  to  market,  rather  than 
those  directly  connected  with  refining,  such  as  cost  of 
power,  are  likely  to  control  the  location  of  a  plant.  Unless 
a  plant  is  built  to  refine  a  large  tonnage  from  a  single  group 
of  mines,  care  must  be  taken  when  determining  upon  its 
location  to  consider  breadth  of  both  copper  supply  and 
market.  The  result  has  been  to  group  a  number  of  the 
American  refineries  around  the  port  of  New  York,  because 
at  that  point  rail  shipments  from  Western  Smelters  and 
ocean  arrivals  fom  South  America  conveniently  join  steam- 
ship deliveries  to  Europe  and  railroads  to  New  England 
mills.  One  of  the  advantages  of  having  a  refinery  close 
to  an  Atlantic  port  is  that  orders  can  be  promptly  filled  in 
the  particular  shape  specified,  enabling  the  European 
customer  to  play  close  to  a  fluctuating  market  which  at 
times  would  result  in  a  preferential  sale.  Another  objection 

183 


184 


COPPER  REFINING 


to  locating  a  refinery  at  a  far  distant  mine  is  that  a  long 
time. elapses  between  the  production  of  a  wirebar  and  its 
actual  rolling,  so  that  there  is  no  prompt  independent  check 
upon  the  quality  of  product  the  plant  is  turning  out.  Aside 
from  these  considerations,  the  availability  of  a  suitable  coal 
for  furnace  work  and  the  existence  of  a  good  boiler  water 
together  with  ample  water  for  condensing  must  be  looked 
into. 

Capacity. — In  the  same  way  external  considerations  may 
justify  the  operation  of  a  plant  of  a  capacity  far  below  the 
economic  size,  but  the  latter  is  a  point  of  universal  interest. 


2000 


8000 


4000 
Short  Tons  per  Month 

FIG.  38. — Variation  of  cost  with  size  of  plant. 


10000 


In  these  days  of  economic  disturbance  it  is  very  hard  to 
state  anything  in  terms  of  absolute  cost,  but  a  relative 
diagram  may  be  made,  showing  with  reasonable  reliability 
the  relation  between  plant  capacity  and  operating  cost. 
This  has  been  attempted  on  a  percentage  basis  in  figure  38, 
where  it  is  shown  that  not  much  is  to  be  gained,  insofar  as 
cost  is  concerned,  by  making  a  plant  of  larger  capacity 
than  200,000,000  Ib.  of  copper  per  year.  Even  this  figure, 
however,  is  beyond  the  reach  of  any  but  the  largest  mining 
groups  so  that  there  will  always  be  a  large  custom  refining 
business. 


ELEMENTS  OF  DESIGN  185 

Tank  House. — Having  decided  upon  the  size  of  plant  to 
be  constructed,  the  next  problem  is  the  choice  of  the  current 
density.  This  question  has  already  been  discussed  at 
length  in  Chapter  IV,  and  in  turn  depends  upon  several 
primary  factors  such  as  cost  of  power,  silver  values,  etc. 
It  has  also  to  be  remembered  that  the  plant  will  earn  more 
interest  on  its  investment  when  operated  at  a  little  higher 
density  than  that  giving  the  lowest  operating  cost,  as 
shown  on  page  69.  On  the  other  hand,  it  is  generally  wise 
to  leave  this  margin  for  the  growth  of  the  business. 

The  density  having  been  settled  upon,  tank  dimensions 
have  next  to  be  considered.  In  the  earlier  plants  electrodes 
about  2  ft.  X  3  ft.,  twenty  pairs  to  the  tank,  were  used  with 
densities  around  15  amp.  per  square  foot.  This  gave 
3,600  amp.  on  a  circuit  which  was  in  those  days  a  heavy 
amperage  to  rectify  on  a  single  commutator.  To-day  30 
pairs  of  electrodes  are  used  without  material  lowering  of 
the  current  efficiency  and  in  one  plant  electrodes  4  ft.  square 
are  in  successful  use,  although  3  ft.  square  is  a  more  popular 
size  where  soluble  anodes  are  employed.  With  30  pairs  of 
3  ft.  electrodes  and  the  average  density  of  20  amp.  per 
square  foot,  the  current  on  a  circuit  becomes  10,800  amp., 
still  well  within  the  limits  of  the  modern  commutator.  At 
an  overall  current  and  time  efficiency  of  85  per  cent,  this 
current  will  produce  17,258  Ib.  of  copper  per  tank  per 
month,  or  575  Ib.  per  pair  of  electrodes. 

The  spacing  center  to  center  of  anodes  will  lie  between  4 
and  4>£  in.,  depending  upon  the  purity  of  the  bullion 
treated  and  the  age  of  electrodes  adopted.  In  general  the 
column  of  liquor  between  anode  and  cathode  should  be 
about  \y±  in.  across.  General  practice  as  to  the  age  ratios 
'of  electrodes  is  indicated  in  Fig.  39.  At  the  lower  densities 
cathodes  and  anodes  are  of  identical  age  wfcile  as  the  density 
rises  increasing  numbers  of.  sets  of  cathodes  are  drawn 
during  the  life  of  one  set  of  anodes.  The  weight  of  the 
anode  can  be  obtained  by  multiplying  the  expected  output 
per  tank-day  per  pair  of  electrodes  by  the  days  life  deter- 
mined upon  and  adding  15  per  cent  to  this  figure  to 


186 


COPPER  REFINING 


allow  for  anode  scrap.  From  the  area  and  weight  (allowing 
something  for  porosity)  the  thickness  may  be  calculated 
and  this  plus  2%  in.  will  give  the  anode  spacing. 

Having  settled  upon  the  output  per  tank-day,  the  total 
number  of  tanks  required  is  at  once  obtained  and  the  next 
question  to  be  considered  is  the  number  of  independent 
circuits  required.  In  general  the  voltage  per  tank  will  be 
somewhere  btween  0.30  and  0.45  volt.  This  figure  depends 
upon  many  factors,  the  chief  being  the  current  density. 


50 


0,20 


10 


12346678 

FIG.  39. — Crops  of  cathodes  during  life  of  anode. 


An  average  value  in  practice  would  be  0.35  volt.  To  this 
must  be  added  a  sufficient  allowance  for  liberator  and 
stripper  tanks.  The  liberators  require  about  2.2  volts 
when  on  normal  work  and  2.5  volts  when  used  for  complete 
removal  of  copper.  Ordinary  work  would  call  for  2  per 
cent  of  the  former  and  none  of  the  latter,  but  this  depends 
upon  the  analysis  of  the  anodes  to  be  handled.  Stripper 
tanks  are  operated  on  a  somewhat  wider  spacing  than 
regular  tanks  and  this  means  fewer  pairs  of  electrodes, 
higher  current  density  and  higher  voltage — say  0.50  volt 
as  compared  witti  0.35.  If  we  take  80  per  cent  of  the  num- 
ber of  electrodes  per  tank,  24-hr,  sheets,  allow  another  20 
per  cent  for  loops  and  bad  production  and  10-day  cathodes, 
not  forgetting  that  one  stripper  blank  yields  two  starting 
sheets,  each  regular  tank  will  demand  the  equivalent  of 


ELEMENTS  OF  DESIGN  187 

3  sheets  a  day  while  each  stripper  tank  will  produce  38 
sheets  a  day. 

Power  House. — The  general  power  problem  was  reviewed 
in  the  last  chapter.  To  this  may  be  added,  however,  some 
notes  on  electrolytic  switchboards. 

Heavy  direct  currents  are  satisfactorily  handled  by  multi- 
blade  knife  switches  with  a  duty  of  50  amp.  per  square- 
inch  of  sliding  contact.  Circuits  are  neither  opened  nor 
closed  under  load  but  a  circuit  breaker  with  a  no-voltage 
release  should  be  placed  on  each  circuit,  not  as  a  protection 
to  the  generator,  the  armature  resistance  of  which  is  too 
low  to  "  short  circuit,"  but  to  protect  men  and  apparatus 
in  the  tank  house.  It  is  not  uncommon  to  have  an  open 
circuit  due  to  the  crane  drawing  cathodes  by  mistake  from 
a  tank  which  has  not  been  cut  out  and  heavy  arcing  may 
result.  A  bar  or  tool  laid  across  lines  where  any  con- 
siderable effective  voltage  is  present  is  promptly  "burned 
out."  An  automatic  engine  stop  should  also  be  provided 
in  the  case  of  reciprocating  engines  as  with  a  sudden  release 
of  load  due  to  the  opening  of  the  circuit  breaker  there  is 
always  a  chance  that  the  governor  will  act  sluggishly,  as  the 
steady  full  load  day  after  day  gives  it  but  little  exercise. 

Recording  and  integrating  instruments  are  not  of  great 
value  on  the  electrolytic  switchboard,  beyond  giving  a 
check  on  the  employes,  as  hourly  readings  of  the  indicating 
instruments  are  ample  for  all  records. 

Several  precautions  are  necessary  in  setting  instruments 
on  the  board.1 

With  voltmeters  there  is  little  trouble,  though  the  same 
precautions  must  be  taken  with  regard  to  the  effect  of  stray 
field  as  stated  below  for  ammeters.  The  very  heavy  cur- 
rents met  with  in  electrolytic  work  accentuate  some  causes 
of  error  which  are  ordinarily  negligible  in  switchboard 
ammeters,  and  if  a  limiting  error  of  1  per  cent  is  to  be  at- 
tained, distribution  of  the  current  in  the  shunt,  thermo- 
electric effects,  switchboard  temperature,  the  magnetic 
effect  of  stray  field  on  the  instrument  and  possible  dynamo 

1  Addicks,  Trans.  Am.  Electrochemical  Society,  vol.  8,  p.  239. 


188  COPPER  REFINING 

action  of  the  leads,  must  be  accounted  for.  One  per  cent 
error  may  be  taken  as  the  limit  of  accuracy  of  the  best 
types  of  switchboard  ammeters  for  heavy  direct  current. 

Care  should  be  taken  to  place  both  shunt  and  instrument 
in  suitable  positions  when  designing  the  switchboard.  In 
the  first  place  the  shunt  should  be  obtained  from  the 
maker  with  short  bars  inserted  in  the  slots  usually  provided 
in  the  shunt  terminals,  these  bars  in  turn  to  be  bolted  to 
the  bus  bar  carrying  the  current  to  be  measured.  In  this 
way  the  shunt  is  calibrated  with  the  same  distribution  of 
current  through  the  various  leaves,  as  will  be  the  case  in 
after  use.  It  is  practically  impossible  to  disconnect  a  high 
capacity  shunt  down  to  the  slotted  terminals  and  put  it 
back  in  place  again  without  altering  the  calibration  more 
or  less,  due  to  change  in  the  distribution  of  the  contact  re- 
sistance between  main  bar  and  shunt  terminals,  and  con- 
sequent slight  change  in  the  amount  of  current  carried  by 
the  different  leaves  of  alloy.  This,  in  turn,  slightly  affects 
the  ratio  between  drop  of  potential  between  binding  posts 
and  total  current  carried.  These  changes  in  resistance  are 
very  small,  but  the  total  resistance  of  a  good  5,000  amp. 
shunt  is  but  about  0.00001  ohm.  There  is  no  way  to  detect 
the  error  introduced  in  this  way  except  by  calibration  in 
place. 

Unequal  heating  of  the  two  shunt  terminals  will  intro- 
duce a  constant  error  from  the  thermo-electric  junctions 
formed  where  the  alloy  leaves  are  soldered  on  to  the  copper 
blocks.  This  may  easily  amount  to  1  per  cent  of  the  full 
scale  reading,  may  be  either  positive  or  negative,  and  will 
show  as  a  zero  error  on  shutting  down  the  circuit  which 
will  gradually  disappear  as  the  shunt  cools  off.  Sometimes 
compensating  devices  are  applied  to  avoid  thermoelectric 
errors,  but  if  the  shunt  is  so  placed  in  the  bus  bar  that  the 
facilities  for  conducting  away  heat  are  approximately  equal 
on  both  sides,  there  will  be  no  appreciable  error,  and  such 
devices  should  be  unnecessary.  The  alloy  leaves  of  a  shunt 
would  be  red-hot  at  full  load  were  it  not  for  the  heat  dissi- 
pating qualities  of  the  large  copper  terminals.  There 


ELEMENTS  OF  DESIGN  189 

should  be  practically  equal  radiating  surfaces  of  bus  bar 
for  three  or  four  feet  each  side  of  the  shunt.  If  placed 
close  to  a  switch  on  one  side,  there  is  sure  to  be  a  thermo- 
electric error  in  the  shunt,  as  high  capacity  switches  frequently 
run  hot.  The  best  plan  is  not  to  attempt  to  place  the  shunt 
back  of  the  switchboard  at  all. 

The  temperature  at  which  a  shunt  runs  will  depend  upon 
how  heavily  it  is  loaded  and  the  opportunity  afforded  to 
dissipate  the  heat  generated.  The  temperature  indicated 
by  a  thermometer  and  well  placed  upon  one  of  the  terminal 
blocks  should  not  be  allowed  to  exceed  200°F.,  owing  to 
the  danger  of  starting  the  solder  and  consequent  failure. 
The  temperature  coefficient  of  the  alloy  used  in  the  resist- 
ance leaves  is  generally  so  low  that  no  correction  to  the 
instrument  reading  need  be  applied  for  the  heating  of  the 
shunt. 

A  sensitive  millivolt  meter  is  used  to  measure  the  drop. 
The  condition  for  maximum  sensitiveness  would  require 
that  all  of  the  resistance  of  the  instrument  be  in  the  arma- 
ture of  the  d'Arsonval  type  so  generally  used.  And  in 
order  that  a  given  resistance  may  afford  as  many  ampere 
turns  as  possible  without  making  too  heavy  a  bobbin,  the 
material  must  be  one  low  in  specific  resistance,  such  as 
copper.  The  use  of  copper  means  a  high  temperature  co- 
efficient, and  in  order  to  keep  within  1  per  cent  on  large 
switchboard  ammeters  it  is  always  necessary  to  correct  for 
instrument  temperature.  The  largest  size  of  instrument 
—independent  of  capacity,  as  change  of  size  for  current 
rating  is  made  in  shunt  arid  not  in  instrument — has  a  tem- 
perature coefficient  corresponding  to  that  of  pure  copper,  or 
0.24  per  cent  per  degree  Fahrenheit.  The  smaller  types  are 
lower  in  temperature  coefficient,  as  the  lighter  movement 
gives  the  instrument  designer  more  leeway.  In  high  re- 
sistance instruments,  such  as  voltmeters  there  is  a  series 
coil,  made  of  low  temperature  coefficient  alloy,  which 
forms  so  large  a  proportion  of  the  total  instrument  resist- 
ance that  the  coefficient  of  the  whole  is  negligible. 

One  of  the  most  important  sources  of  error  is  the  magnetic 


190  COPPER  REFINING 

effect  of  stray  field  in  the  immediate  vicinity  of  a  bar 
carrying  several  thousand  amperes.  Instruments  should 
never  be  placed  within  two  feet  of  such  bars.  The  iron 
shield  usually  provided  is  not  only  likely  to  be  faulty 
as  a  filter  for  lines  of  force  of  such  density,  but  consequent 
poles  are  formed  in  the  case  itself. 

The  magnet  of  an  unscreened  Weston  portable  instru- 
ment may  be  permanently  altered  by  placing  the  meter  too 
close  to  such  a  bar,  and  this  is  very  easy  to  do,  especially 
in  the  tank  house.  Even  at  a  distance  of  several  feet  the 
readings  of  a  portable  instrument  are  thrown  out  a  division 
or  two  on  the  scale  and  where  accuracy  is  desired  the 
average  of  two  readings  with  the  instrument  turned  180 
deg.  should  be  used. 

The  resistance  of  the  leads  on  a  switchboard  ammeter  is 
generally  some  5  per  cent  of  that  of  the  instrument,  and 
the  screw  connections  must  be  kept  clean.  If  it  is  found 
necessary  to  lengthen  the  leads,  care  must  be  taken  that  the 
resistance  is  kept  constant  by  a  proportionate  increase  in 
size  of  wire  used.  The  leads  should  be  twisted  and  kept 
from  swinging  loosely  in  the  presence  of  stray  field.  Other- 
wise there  is  a  dynamo  action  which  the  millivoltmeter  is 
sensitive  enough  to  indicate  by  a  pulsating  needle. 

When  several  high  capacity  ammeters  are  in  use,  it  is 
advisable  to  have  them  interchangeable.  The  various 
instruments  may  then  be  checked  at  any  time  by  con- 
necting them  to  the  same  shunt.  Connections  should  also 
be  arranged,  if  possible,  to  put  the  shunts  in  series,  when 
they  too  may  be  checked. 

The  question  of  proper  section  for  bus  bars  has  been 
discussed  on  page  43.  No  insulation  is  used  and  faced 
metal  to  metal  joints  without  amalgam,  tightly  bolted 
together  are  satisfactory  at  200  amperes  per  sq.  in.  contact 
surface.  Either  rolled  or  cast  bars  may  be  employed. 
The  bars  from  the  switchboard  to  the  tanks  are  best 
carried  overhead  and  the  positive  and  negative  leads 
kept  some  distance  apart. 


ELEMENTS  OF  DESIGN 


191 


Furnaces. — For  smooth  operating  there  should  be  at 
least  six  furnaces,  three  anode  and  three  wirebar  units, 
two  of  each  being  in  service  and  one  out  for  repairs.  For  a 
capacity  of  100,000  tons  a  year,  and  allowing  for  Sundays 
and  holidays  out,  we  should  have  600  charges  a  year  of 
333,000  Ib.  each  plus  scrap,  etc.  which  is  a  furnace  of 
fair  size,  although  it  is  possible  to  build  them  very  much 
larger. 


50 


75       100      125       150 

Short  Tons  .per  Charge 
FIG.  40. — Fuel  ratio  of  furnaces. 


175      200 


The  limit  of  size  has  been  the  ability  to  charge,  refine 
and  cast  the  charge  in  twenty-four  hours.  With  hand 
charging  and  ladling,  using  the  same  men  for  both  opera- 
tions, 60,000  Ib.  is  a  large  furnace.  If  a  fresh  crew  is  used 
for  charging  and  as  many  ladlers  employed  as  can  be 
accommodated  at  the  ladle  door,  100,000  Ib.  can  be  reached. 
With  hand  charging  but  mechanical  ladling  300,000  Ib. 
is  possible,  using  three  charging  doors.  With  full  mechan- 
ical charging  and  ladling  500,000  Ib.  is  easily  reached  and 
using  two  charging  and  ladling  machines  even  a  larger 
capacity  can  be  attained. 


192  COPPER  REFINING 

The  fuel  economy  increases  rapidly  with  the  size  of 
furnace  as  indicated  by  the  curve  in  Fig.  40.  Labor 
charges  also  fall  up  to  the  point  of  mere  duplication  of 
machine  units.  Repairs  fall  at  first  but  a  very  large 
furnace  is  likely  to  incur  heavier  repairs  than  one  somewhat 
smaller.  This  is  especially  true  where  too  great  a  span 
of  roof  is  attempted.  A  refining  furnace  is  subjected  to 
much  greater  and  more  rapid  temperature  changes 
than  an  ore  reverberatory. 

Small  silica  furnaces  produce  excessive  slag  as  they  run 
hot  and  have  an  undue  proportion  of  wall  exposure.  A 
40,000  Ib.  silica  furnace  will  form  4  or  5  per  cent  of 
slag  on  a  wirebar  charge,  whereas  a  similar  furnace  of 
150,000  Ib.  or  over  to  the  charge  can  be  held  down  to  half 
that  amount.  With  basic  side  walls  there  is  a  substantial 
decrease  in  these  figures  and  with  a  full  basic  furnace 
operating  on  a  charge  of  good  quality  the  slag  made  can 
be  held  down  to  0.5  per  cent  or  bettei. 

Altogether  a  furnace  casting  a  charge  of  300,000  Ib.  or 
more  may  be  considered  a  thoroughly  economical  unit. 

Silver  Refinery.— A  100,000  ton  plant  will  call  for  a 
silver  refining  department  of  reasonable  size.  At  1.25 
per  cent  slimes  there  would  be  four  tons  of  slimes  a  day  to 
treat  calling  for  a  dore  reverberatory  of  about  20  tons 
capacity,  just  the  size  of  the  first  copper  furnaces.  It  is 
desirable  to  use  but  a  single  furnace  in  order  to  hold  at  a 
minimum  the  large  bottom  absorption  of  values.  A  spare 
furnace  may  be  held  in  reserve  and  in  case  of  accident  or  at 
inventory  time  substituted,  the  first  unit  being  completely 
dismantled  at  that  time  and  its  silver  content  recovered. 

A  very  small  plant  can  best  sell  its  slimes  to  a  lead  re- 
finery but  owing  to  the  difficulty  in  getting  a  fair  sample 
of  such  rich  material  this  is  not  desirable  when  the  tonnage 
is  of  any  size.  Small  plants  would  use  a  cupel  furnace 
and  sulphuric  acid  parting  while  reverberatory  smelting  fol- 
lowed by  electrolytic  parting  is  generally  employed  in 
large  ones. 

Where  the  purifying  system  affords  a  suitable  outlet  for 


ELEMENTS  OF  DESIGN  193 

accumulations  of  sodium  salts  in  the  electrolyte  niter 
may  be  used  to  remove  the  copper  from  the  slimes.  Where 
closed  cycle  operation  is  undertaken  the  copper  is  generally 
oxidized  by  roasting.  Either  method  gives  good  metallur- 
gical results. 

Purifying. — The  metallurgical  principles  upon  which  the 
choice  of  a  purilying  system  must  be  based  were  taken 
up  in  Chapter  VI. 

Where  insoluble  anode  tanks  are  employed  for  the 
complete  removal  of  the  copper  from  batches  of  electrolyte, 
ample  provision  must  be  made  for  the  very  low  current 
efficiency  of  these  tanks.  Should  a  bluestone  plant  be 
employed,  the  design  will  be  based  upon  the  quantity  of 
shipping  salt  to  be  produced.  The  crystallizing  tank  space 
employed  in  different  plants  varies  considerably.  A  fair 
average  figure,  making  allowance  for  the  recrystallization 
of  fines,  is  58  cu.  ft.  per  short  ton  per  month  of  marketable 
bluestone.  When  estimating  the  proportion  of  impurities 
removed  from  the  electrolyte  it  must  be  remembered  that 
a  large  part  of  the  copper  in  the  bluestone  comes  from 
the  shot  towers  and  not  from  the  electrolyte.  For  each 
1000  cu.  ft.  of  crystallizing  space,  allowance  should  be  made 
for  say  175  cu.  ft.  shot  towers,  375  cu.  ft.  shot  tower  re- 
ceiving tanks,  225  cu.  ft.  slimes  settling  tanks,  250  cu.  ft. 
boiling  tanks,  and  275  cu.  ft.  iron  cementation  tanks. 
These  ratios,  however,  may  be  varied  between  rather  wide 
limits  to  suit  individual  cases  differing  in  analysis  of  elec- 
trolyte, character  of  shot,  climate,  etc. 

General. — Two  main  principles  should  be  kept  in 
mind  in  the  program  of  operation  of  any  plant.  The  first  is 
that  all  the  impurities  possible  should  be  eliminated  at  the 
anode  furnaces,  using  basic  furnaces  if  necessary.  It  is  much 
more  costly  to  throw  this  burden  upon  the  tank  house  and 
silver  refinery. 

The  second  is  to  do  everything  possible  to  facilitate  uni- 
formity of  operating  conditions.  As  a  general  rule  anodes 
of  constant  composition,  a  uniform  current  density  and  a 
single  electrolyte  should  be  use'd  throughout  the  tank  house. 

13 


CHAPTER  XIII 
APPLICATION  TO  OTHER  FIELDS 

Electrolytic  refining  has  attained  its  chief  application  and 
highest  development  in  the  metallurgy  of  copper.  The 
very  dynamo,  the  invention  of  which  created  the  large-scale 
demand  for  high-conductivity  copper,  gave  the  means  of 
producing  it  by  electro-deposition,  and  electrolytic  refining 
on  a  commercial  scale  was  one  of  the  first  electrical  indus- 
tries to  be  developed,  while  it  remains  today  by  far  the 
largest  application  of  the  electrolysis  of  an  aqueous  elec- 
trolyte. 

The  experience  gained  in  this  field  has  found  wide  appli- 
cation in  the  refining  of  other  metals,  such  as  silver,  gold, 
lead,  bismuth,  tin,  nickel,  iron  and  zinc,  and  in  the  recovery 
direct  from  the  ore  by  leaching  of  copper,  silver,  gold  and 
zinc.  The  successful  application  of  electrolysis  to  these 
different  fields  requires  a  rebalancing  of  the  various  factors 
discussed  in  the  foregoing  chapters,  the  relative  values  of 
which  are  often  greatly  altered  by  a  change  of  metal.  In 
any  given  case  we  must  take  into  account  (1)  competition 
from  other  processes,  (2)  acid  radical  to  be  employed,  (3) 
temperature  of  electrolyte,  (4)  character  of  deposit,  (5) 
resolution  at  the  cathode,  and  (6)  depolarization  at  the 
anode. 

Competition. — Electrolytic  copper  refining  has  no  ef- 
fective competitor  in  its  own  field.  Fire  refining  makes  a 
low  conductivity  product  unless  furnished  with  very 
pure  raw  material  such  as  selected  Lake  "mineral,"  and 
the  various  selective  methods  of  reverberatory  treatment 
yield  but  a  partial  recovery  of  gold  and  silver. 

In  the  leaching  field  electrolysis  has  to  meet  several 
competitive  methods  of  reduction,  notably  precipitation 
upon  iron.  While  a  free  acid  equivalent  by  electrolyzing 

194 


APPLICATION  TO  OTHER  FIELDS  195 

is  returned  to  the  cycle,  there  are  usually  serious  handicaps 
imposed  by  large  quantities  of  impurities  which  cause 
various  troubles  in  the  cells  and  which  progressively 
accumulate  unless  some  outlet  is  provided.  Even  where 
electrolysis  is  indicated,  therefore,  an  iron  precipitation 
plant  is  usually  required  as  an  adjunct,  from  10  to  30  per 
cent  of  the  copper  being  recovered  as  cement. 

Electrolytic  silver  has  to  compete  with  the  older  sulphuric 
acid  method  of  parting.  The  latter  is  easy  to  operate 
and  ties  up  less  silver,  but  it  will  not  make  a  product  so 
low  in  gold  as  the  electrolytic.  It  is  still  indicated,  however 
for  small  plants  and  for  those  which  are  used  spasmodically. 

In  the  case  of  gold  refining  the  question  is  almost  wholly 
dependent  upon  the  amount  of  platinum  and  associated 
metals  present.  If  the  original  gold  is  free  from  this  metal 
group,  there  is  no  object  in  tieing  up  the  gold  several  days 
for  electrolysis  and  further  increasing  the  opportunities  for 
theft.  When  but  very  small  quantities  of  platinum  and 
palladium  are  present,  they  may  be  satisfactorily  collected 
by  wet  methods  instead  of  by  electrolysis  at  about  equal 
expense.  When  the  qualities  are  larger  electrolysis  is 
indicated. 

Electrolytic  lead  has  to  compete  with  the  very  efficient 
Parkes  process  of  refining.  When  all  factors  are  considered, 
the  justification  for  electrolysis  appears  to  rest  almost 
entirely  upon  the  quantity  of  bismuth  present  in  the  bullion 
to  be  treated.  This  objectionable  impurity  is  not 
satisfactorily  removed  by  the  Parkes  process  and  special 
treatments  such  as  Pattisonizing  are  expensive,  while  the 
electrolytic  method  readily  separates  the  bismuth  from  the 
lead  and  converts  it  into  a  marketable  byproduct. 

Nickel,  iron  and  zinc  may  be  readily  refined  by  elec- 
trolysis, but  in  ordinary  times  the  value  of  the  refined 
product,  except  in  a  very  limited  quantity,  is  not  suffi- 
ciently above  that  of  the  crude  to  pay  for  the  refining. 
Tin  offers  a  special  field  for  impure  ores  which  are 
difficult  to  handle  by  fire  processes. 

In  leaching,  both  nickel  and  zinc  have  strong  competition 


196  COPPER  REFINING 

from  chemical  and  pyrometallurgical  processes  and  this 
whole  field  is  still  in  a  state  of  flux. 

In  general,  we  can  say  that  electrolysis  enjoys  no  such 
absolute  monopoly  in  other  fields  as  it  does  in  copper  refin- 
ing and  that  each  proposed  application  must,  therefore, 
receive  full  consideration  upon  its  merits. 

Acid  Radical. — While  several  salts  of  copper  are  suitable 
for  electrolytic  treatment,  the  sulphate  possesses  so  many 
practical  advantages  that  it  is  universally  employed.  The 
first  requirement  for  an  electrolyte  is  that  the  salt  of  the 
metal  employed  shall  be  readily  soluble.  The  second  is 
that  the  material  which  it  is  desired  to  collect  in  the 
slime  shall  be  insoluble  in  it.  The  third  is  that  the  cathode 
should  not  be  redissolved  by  the  electrolyte.  Copper 
sulphate  is  a  readily  soluble  salt  and  copper,  gold  and  silver 
are  all  practically  unattacked  by  dilute  sulphuric  acid. 

When  electrolyzing  silver  a  nitrate  electrolyte  is  em- 
ployed, as  silver  sulphate  is  not  a  readily  soluble  salt.  The 
amount  of  free  acid  carried  is  necessarily  low,  as  nitric 
acid  is  expensive  and  decomposes  under  the  conditions  of 
electrolysis.  Copper  concentrates  in  the  electrolyte  and  is 
removed  by  withdrawals,  while  gold  remains  unattacked 
in  the  slimes. 

In  gold  refining  a  chloride  electrolyte  is  employed. 
Gold  chloride  is  a  soluble  salt,  platinum  concentrates  in 
the  electrolyte,  and  although  silver,  being  less  noble  than 
gold,  dissolves  at  the  anode,  after  saturation  it  precipitates 
as  a  secondary  slime  of  insoluble  chloride. 

For  lead,  chloride  and  acetate  were  first  tried,  but  not 
until  Betts  developed  the  fluosilicate  electrolyte  was  the 
electrolytic  refining  of  lead  made  a  commercial  success. 

It  will,  therefore,  be  appreciated  that  an  entirely  new 
set  of  chemical  conditions  as  regards  behavior  of  impuri- 
ties, etc.,  is  encountered  with  every  metal  refined. 

Temperature  of  Electrolyte. — When  it  is  customary  to 
heat  copper  electrolytes  externally,  quite  opposite  condi- 
tions obtain  in  the  case  of  silver,  lead  and  zinc.  Nitric 
and  hydrofluosilicic  acid  show  increasing  decomposition 


APPLICATION  TO  OTHER  FIELDS  197 

losses  with  rise  in  temperature,  asphalted  tanks  do  not  stand 
hot  liquors  and  zinc  resolution  losses  must  be  limited.  In 
the  case  of  silver,  the  current  may  have  to  be  limited  to 
control  the  temperature;  for  lead,  extraneous  heating  is 
abandoned,  and  with  zinc,  where  the  specific  resistance  of 
the  electrolyte  is  high,  cooling  systems  carrying  circulating 
water  must  be  employed. 

Character  of  Deposit. — At  moderate  densities  copper 
readily  gives  a  good  adherent  deposit.  The  same  may  be 
said  of  gold.  Silver  gives  normally  a  loose  crystalline 
deposit.  Nickel  is  nodular,  while  lead  and  zinc  tree  very 
badly.  By  the  proper  use  of  addition  agents  all  of  these 
deposits  may  be  made  smooth  and  coherent.  In  fact  it 
was  not  until  the  value  of  addition  agents  was  more  or  less 
understood  that  lead  could  be  handled  at  all. 

In  the  early  copper  plants  a  great  deal  of  trouble  was 
had  in  getting  good  starting  sheets,  and  it  was  thought 
that  the  addition  of  ammonium  sulphate  was  a  help,  the 
use  of  double  salts  being  common  practice  in  various 
electroplating  prescriptions.  Then  the  beneficial  effect  of 
the  presence  of  wood  strips  on  the  edges  of  the  starting 
blanks  led  to  the  introduction  of  wood  tea  made  from  shav- 
ings, yielding  a  complex  organic  addition  agent.  Finally 
the  remarkable  results  following  the  addition  of  gelatine 
were  worked  out.  The  irregularities  of  early  practice 
were  due  to  overdosing  and  a  failure  to  appreciate 
the  fleeting  effect  of  a  single  dose.  Today  a  few  pounds  of 
common  glue  dissolved  in  water  and  added  regularly 
throughout  the  twenty-four  hours,  together  with  a  can  of 
ordinary  engine  oil,  will  control  the  deposit  in  the  tank 
house  of  a  great  copper  refinery,  permitting  the  use  of 
higher  densities,  closer  spacing  and  greater  cathode  age. 

For  silver,  while  adherent  deposits  can  be  made  in 
similar  fashion,  it  is  customary  to  collect  the  silver  crystals 
as  such,  either  brushed  to  the  floor  of  the  tank  from  vertical 
cathodes  by  mechanical  scrapers  as  in  the  Moebius  system 
or  shoveled  by  hand  from  the  horizontal  carbon  cathode 


198  COPPER  REFINING 

of  the  Thum  cell.  These  crystals  are  readily  washed  free 
of  electrolyte  and  fed  to  retorts  for  melting. 

Lead  may  be  beautifully  controlled  by  addition  agents. 
The  exact  amount  to  be  added  varies  from  time  to  time 
and  is  determined  by  experiment  and  careful  watching 
of  the  fresh  deposit  on  special  strips  hung  for  the  purpose. 

Zinc  is  generally  plated  in  coherent  form  on  an  aluminum 
cathode,  from  which  it  is  peeled  every  forty-eight  hours. 
There  is  no  reason  why  zinc  starting  sheets  could  not  be 
used  as  in  the  case  of  copper,  and  in  some  ways  the  avoid- 
ance of  a  zinc-aluminum  couple  would  be  of  advantage. 
In  the  development  of  electrolytic  zinc  practice,  however, 
there  has  been  great  difficulty  in  controlling  resolution  at 
the  cathode,  and  the  great  advantage  of  using  aluminum 
sheets  lies  in  the  fact  that  strong  conosion  can  do  no  more 
than  leave  the  starting  sheets  bare,  whereas  were  they  of 
zinc  they  might  be  completely  eaten  through,  causing 
collapse,  open  circuits,  general  demoralization  and  shut- 
down of  the  plant. 

Resolution  at  Cathode. — The  cathode  is  never  absolutely 
insoluble  in  the  electrolyte  used  and  a  small  amount  of 
reoxidation  is  always  in  process.  In  the  case  of  copper  with 
normal  electrolytes  it  is  a  minor  matter — perhaps  2  per 
cent.  Should  we  allow  an  accumulation  to  take  place  of 
salts  of  manganese,  iron  or  any  other  metal  capable  of 
alternate  oxidation  and  reduction  at  anode  and  cathode 
with  change  of  valence,  a  very  serious  condition  may  arise. 

In  these  days  copper  anodes  are  so  highly  refined  that 
the  amount  of  iron  contained  in  the  electrolyte  is  quite 
negligible,  but  when  electrolyzing  liquors  arising  from 
the  leaching  of  ore  a  very  different  situation  exists.  Such 
liquors  commonly  contain  large  quantities  of  ferrous  sul- 
phate and  a  certain  proportion  of  ferric  sulphate.  Oxida- 
tion at  the  anode  tends  to  increase  the  latter  at  the  expense 
of  the  former.1  Experiments  show  that  0.25  per  cent  of 
iron  as  ferric  sulphate  in  an  electrolyte  sufficiently  agitated 
will  corrode  cathode  copper  at  a  rate  which  will  require  a 

1  See  Addicks,  Trans.  Am.  Electroch.,  Soc.,  vol.  28,  p.  87. 


APPLICATION  TO  OTHER  FIELDS  199 

current  corresponding  to  about  8  amperes  per  sq.  ft.  merely 
to  replace  the  loss.  The  proper  control  of  ferric  sulphate  is, 
therefore,  the  key  to  success  in  electrolyzing  copper  leaching 
liquors. 

Another  interesting  example  of  resolution  is  in  the 
case  of  zinc — in  fact  this  is  the  controlling  factor  in  zinc 
electrolysis.  In  the  electrolysis  of  zinc  sulphate  using  an 
insoluble  anode  the  content  of  free  sulphuric  acid  gradually 
increases  and  unless  the  cathode  is  absolutely  pure  zinc 
resolution  is  very  active.  Perfectly  pure  zinc  is  so  nearly 
insoluble  in  sulphuric  acid  that  it  is  quite  difficult  to  get  it  in 
solution  for  analysis.  But  if  a  nodule  of  cathode  zinc  is 
immersed  in  dilute  sulphuric  acid  in  a  test  tube  and  the 
merest  trace  of  almost  any  impurity  is  added,  a  vigorous 
evolution  of  hydrogen  will  start  at  once.  This  is  probably 
due  to  galvanic  action,  and  the  whole  secret  of  successful 
zinc  electrolysis  lies  in  the  passivity  of  pure  zinc  in  sul- 
phuric acid.  This  calls  for  a  degree  of  purity  in  the  elec- 
trolyte which  would  be  quite  uncommercial  were  it  not  for 
the  fact  that  zinc  dust  readily  throws  down  most  of  the 
impurities  to  be  dealt  with  after  the  liquor  from  the  leach- 
ing tanks  has  been  neutralized.  The  last  traces  of  some 
difficult  elements,  such  as  arsenic,  are  removed  by  adsorp- 
tion upon  freshly  made  ferric  hydrate. 

Depolarization  at  the  Anode. — Anode  efficiency  plays  a 
relatively  small  part  in  normal  copper  refining;  the  propor- 
tion of  impurities  present  in  the  anode  is  small  and  the 
useful  anode  efficiency  very  high.  As  soon  as  a  complex 
or  insoluble  anode  is  substituted,  however,  the  oxidizing 
effect  of  the  current  is  in  part  or  in  whole  employed  in  the 
solution  of  anode  impurities,  the  oxidation  of  suitable 
salts  in  the  electrolyte  or  in  the  decomposition  of  water  with 
escape  of  oxygen  as  such.  This  introduces  several  new 
problems. 

A  moderate  amount  of  an  oxidizable  impurity,  such  as 
nickel,  in  a  copper  anode  gives  the  first  and  simplest  case. 
The  nickel  and  copper  dissolve  proportionately  at  the  anode, 
sharing  the  current.  At  the  cathode,  however,  only 


200  COPPER  REFINING 

copper  is  deposited,  so  that  the  copper  in  the  electrolyte 
is  correspondingly  depleted.  About  2  per  cent  of  the 
copper  deposited  is  restored  by  purely  chemical  solution 
at  the  electrodes  but  beyond  that  figure  soluble  copper 
must  be  added  by  leaching  shot  or  scale,  and  the  accumulat- 
ing nickel  must  be  controlled  by  withdrawals  to  a  byproduct 
plant. 

The  second  difficulty  arises  when  the  quantity  and  nature 
of  the  impurity  in  the  anode  causes  segregation  into  two 
than  the  components  one  of  which  is  more  readily  dissolved 
other.  Then  the  anode  disintegrates  unevenly  and  a 
large  amount  of  scrap  has  to  be  reworked. 

A  third  and  more  serious  condition  is  met  with  when  the 
impurity  is  insoluble,  such  as  lead  or  antimony,  when  the 
anode  will  become  coated  with  a  non-conducting  slime. 
The  voltage  will  then  rise  until  some  free  oxygen  is  gene- 
rated from  the  moisture  underneath.  This  in  turn  will 
burst  through  the  coating  and  the  anode  will  act  normally 
for  a  few  seconds,  when  the  coating  will  again  form.  A 
voltmeter  connected  across  such  a  tank  will  show  a  wildly 
fluctuating  needle,  and  this  condition,  known  colloquially 
as  " crazy  tanks,"  is  fatal  to  good  refining,  entailing 'as 
it  does  excessive  gold  and  silver  losses  in  the  cathode,  high 
power  cost  and  a  heavy  expense  for  purifying  electrolyte. 
The  remedy  lies  in  properly  refining  the  crude  material 
before  casting  the  anodes. 

Finally,  we  have  the  extreme  case  where  the  anode  is 
by  intent  insoluble,  as  where  copper  is  being  recovered 
from  leaching  liquor.  Where  no  depolarizer  is  employed, 
the  voltage  must  of  course  be  sufficiently  high  to  decom- 
pose water,  and  free  oxygen  is  given  off  at  the  anode. 
When  the  liquor  is  virtually  free  from  chlorides  and  nitrates, 
antimonial  lead  is  generally  used  as  the  anode  material; 
when  electrolyzing  zinc,  however,  pure  lead  is  required  in 
order  not  to  poison  the  cathode  with  specks  of  antimony. 
When  the  liquor  is  corrosive,  either  magnetite  or  one  of  the 
ferro-alloys,  generally  ferro-silicon,  is  employed.  None  of 
these  anode  materials  is  wholly  free  from  oxygen  attack. 


APPLICATION  TO  OTHER  FIELDS  201 

Lead  peroxidizes  and  sulphatizes  and  the  iron  alloys  slowly 
dissolve,  so  that  a  certain  replacement  charge  must  be 
reckoned  with. 

Where  an  efficient  depolarizer  is  employed,  any  of  the 
materials  mentioned  above  or  carbon  may  be  employed. 
Graphite  offers  peculiar  advantages  in  that  100  per  cent  of 
efficiency  of  oxidation  of  ferrous  sulphate  or  similar  depolar- 
izer is  readily  obtained  by  its  use.  Lead  does  not  give 
equal  results.  On  the  other  hand,  carbon  itself  will  oxidize 
and  disintegrate  if  not  fully  protected  by  the  depolarizer. 
This  whole  question  of  a  cyclic  oxidization  at  the  anode 
with  subsequent  reduction  at  the  ore  contact  is  yet  in 
but  partially  developed  form,  most  of  the  theoretical 
advantages  being  generally  offset  in  practice  by  the  diffi- 
culties met  with  in  handling  impurities  dissolved  from  the 
ore. 


INDEX 


Addition    agents,     effect    on    con- 
ductivity of,  52 

effect  on  voltage  of,  50 

function  of,  75,  197 

in  electrolytic  parting,  116 
Alumina,  in  dore  furnace  slags,  112 

in  electrolyte,  88 
Aluminum,    effect   on    conductivity 

of,  144 

Amalgamation  plates,  sampling  of, 
160 

treatment  of,  161 
Ammeters,  precision  of,  187 

proper  location  of,  190 
Analysis  of,  anodes,  85,  87 

anode    furnace    slag,    25,     137 

blister  copper,  82,  166 

bluestone,  96 

boneash,  135 

cement  copper,  91.  165 

crude  nickel  salt,  100 

cupel  slag,  111 

dore,  112 

dore  furnace  flue  dust,  112 

dore  furnace  slags,  112 

electrolyte,  88 

electrolytic  nickel,  104 

float  slime,  86 

foreign  pig  copper,  81,  166 

Korean  coins,  166 

mother  liquor  salt,  96 

nickel  salts,  105 

pig  copper,  80,  166 

refining  furnace  slag,   28,    137, 
167 

secondary  pig  copper,  81,  166 

silver  electrolyte,  115 

slimes,  87,  106 

slimes  by  screen,  118,  120 

wirebars,  85 


Annealing,    effect    of,    on  conduc- 
tivity, 146,  147 

when    measuring  conductivity, 

148 
Anodes,  age  of,  63 

converter,  106,  107 

ferro-silicon,  200 

forms  of,  46 

graphite,  201 

impurities  in,  85,  87 

lead,  201 
Anode  scrap,  amount  of,  26 

in  parting,  115 
Anode  slimes,  boiling,  108 

composition  of,  12,  87,  106,  112 

copper  in,  27,  87,  106 

leaching,  110,  121 

oxidation  of,  110,  120 

roasting,  109,  120 

screening,  108,  117 

smelting,  110,  121 

value  of,  108 

Antimony,  effect  of,  on  conductivity, 
144- 

in  anodes,  85,  87 

in  blister  copper,  82,  166 

in  bluestone,  95,  96 

in  dore",  112 

in  dore  furnace  flue  dust,  112 

in  dore  furnace  slags,  112 

in  electrolyte,  88 

in  float  slime,  86 

in  foreign  pig  copper,  81,  166 

in  Korean  coins,  166 

in  mother  liquor  salt,  96 

in  nickel  salts,  105 

in  pig  copper,  80,  166 

in  secondary  pig  copper,  81,  166 

in  slimes,  87,  106,  112 

in  wirebars,  85 

marketed,  as  105 

removal  of,  from  electrolyte,  27 


203 


204 


INDEX 


Arsenic,  effect  of,  on  conductivity,  144 

in  anodes,  85,  87 

in  anode  furnace  slags,  137 

in  blister  copper,  82,  166 

in  bluestone,  95,  96 

in  crude  nickel  salt,  100 

in  dore,  112 

in  dore"  furnace  flue  dust,  112 

in  dore  funace  slags,  112 

in  electrolyte,  88,  89 

in  electrolytic  nickel,  104 

in  float  slime,  86 

in  foreign  pig  copper,  81,  166 

in  iron  tank  cement,  91 

in  Korean  coins,  166 

in  mother  liquor  salt,  96 

in  nickel  salts,  105 

in  pig  copper,  80,  166 

in  secondary  pig  copper,  81,  166 

in  slimes,  87,  106,  112 

in  wirebars,  85,  89 

marketed  as,  105 

removal  of,  by  liberating  tanks, 
97 

removal  of,  from  copper,  165 

removal  of,  from  electrolyte,  27 

uses  of,  101 

Arseniuretted  hydrogen,  97 
Assaying,  combination  assay,  8 

corrected  assays,  9 

cupeling,  8 

electrolytic  assay,  8 

errors  in,  10 

fire  assay,  9  j 

losses  in,  8  | 

methods,  8 

mercury  assay,  9 

splitting  limits  in,  9 

sulphuric  acid  assay,  9 
'  umpire  assays,  10 

B 

Bismuth,  effect  of,  on  conductivity, 

144 

in  anodes,  85 
in  dore,  112 

in  dore  furnace  slags,  112 
in  electrolyte,  88 
in  float  slime,  86 


Bismuth,  in  Korean  coins,  166 

in  secondary  pig  copper,  166 

in  slimes,  87,  106,  112 

in  wirebars,  85 

marketed  as,  105 

uses  of,  101 

Black  copper,   composition   of,   80, 
139,  166 

slags,  11 
Bluestone,  boiling  tanks,  94 

composition  of,  96 

copper  in,  96 

crystallizing  tanks,  94 

crystals,  95 

impurities  in,  95 

manufacture  of,  92 

plant  capacity,  93 

purification  of,  95 

shot  towers,  93 

Boilers,  (see  waste  heat  boilers) 
Boiling  tanks,  capacity  of,  193 

for  crude  nickel  salts,  98 

for  slimes,  108 

operation  of,  .94 

steam  required  for,  181 
Boneash,  135 
Busbars,  (see  conductors) 
By-products,  101 


Cadmium,  effect  of,  on  conductivity, 

144 
Cathodes,  age  of,  64 

conductivity  of,  146 

corrosion  of,  77 

cuprous  chloride  in,  77 

forms  of,  46 

gold  and  silver  losses  in,  13 
Calcium  sulphate,  in  bluestone,  95 

in  electrolyte,  88 
Cementcopper,compositionof,91,165 

from  electrolyte,  91 

iron  consumption  in  producing, 

91 

Charging,  126,  191 
Chlorides,  in  cathodes,  77 

in  electrolyte,  83,  88,  196 
Chrome,  brick  in  furnace  construc- 
tion, 164 


INDEX 


205 


Circulation  of  electrolyte,  effect  of, 

59 

limit  of,  49 

Coal,  consumption  at  furnaces,  172 
Cobalt,   in  electrolytic  nickel,    104 

in  nickel  salts,  105 
in  pig  copper,  166 

in  slimes,  106 

marketed  as,  105 
recovery  of,  168 
Coking,  130 

Concentration,  of  anode  slimes,  120 
Condenser  water,  177 
Conductivity,  effect  of  annealing  on, 
146,  147 

effect  of  hard  drawing  on,   145, 
147 

effect    of    impurities    on,     144 

Matthiessen's  standard  of,  146 

of  cathodes,  146 

of  electrolytic  copper,  143 

of  lake  copper,  143 
Conductors,    contact    resistance    in 
joints,  ofj  45,  190 

corrosion  of,  46 

economical  section  of,  43 
Contacts,  heat  from  48 

lessening,  48 

nature  of,  45 

number  of,  47 

resistance  of,  42,  45 
Copper,  best  selected,  124 

brittleness  in,  151,  152 

conductivity  of,  143 

in  anodes,  85 

in  anode  furnace  slag,  137 

in  black  copper,  166 

in  blister  copper,  82,  166 

in  bluestone,  96 

in  crude  nickel  salt,  100 

in  cupel  slag,  111 

in  dore",  112 

in  dore  furnace  flue  dust,  112 

in  dore  furnace  slag,  112 

in  electrolyte,  88 

in  electrolytic  nickel,  104 

in  float  slime,  86 

in  foreign  pig,  81,  166 

in  Korean  coins,  166 


Copper,  in  mother  liquor  salt,  96 
in  nickel  salts,  105 
in  pig  copper,  80 
in  refining  furnace  slag,  137 
in  secondary  pig  copper,  81,  166 
in  slimes,  27,  87,  112 
in  wirebars,  85 
Lake,  82,  124 
overpoled,  132,  150 
refining  time,  19 
shot,  93 
Corrosion,  at  cathode,  77 

effect  of  ferric  salts  on,  78,  198 
effect  of  temperature  on,  58 
of  conductors,  46 
offset  by  nickel,  59 
Cost,  investment,  184 
labor,  62,  126,  141 
operating,  68,  184 
Cost  of  plant,  relation  to    current 

density,  62 

relation  to  tonnage,  184 
Cottrell  system,  12 
Counter  electro-motive  force,  mag- 
nitude of,  50 
Crystallizing  tanks,  capacity  of,  193 

operation  of,  94 
Current  density,  at  busbar    joints, 

190 

at  switches,  187  ] 
metallurgical  effects  of,  64 
relation  to  age  of  cathodes,  64 
relation  to  cost  of  labor,  62 
relation  to  cost  of  operating,  68 
relation  to  cost  of  plant,  62 
relation  to  cost  of  power,  60 
relation  to  electrode  spacing,  51 
relation  to  metals  tied  up,  66 
Current  efficiency,  attained  in  prac- 
tice, 71 

corrosion  factor  in,  77 
current  leakage  factor  in,  72 
economic,  79 
effect  of  nodules  on,  78 
of  liberating  tanks,  97 
reaction  factor  in,  76 
Current  leakage,  between  electrodes, 

74 
effect  on  current  efficiency,  72 


206 


INDEX 


Current  leakage,  measurement  of,  73 

through  electrolyte,  73 

to  ground,  72 
Cupola  furnace,  charge,  139 

for  smelting  junk,  161 

product,  139 

size,  139 

slag,  11 
Cuprous  salts,  at  anodes,  107 

effect  on  cathodes,  77 

effect  on  current  efficiency,  76 


Electrolyte,  free  acid  in,  52 
growth  of  copper  in,  58 
heating,  54,  57,  180 
quantity  of,  34 
specific  resistance  of,  52 
temperature  coefficient  of,  53 
temperature  of,  53,  54 
Electrostatic  precipitation,  12 
Engines,  (see  steam,  gas,  etc.) 
Evaporators,  (see  boiling  tanks) 


Depolarization,  199 
Design,  furnace,  191 

general,  183,  193 

power  house,  187 

silver  refinery,  192 

sulphate  building,  193 

switchboard,  187 

tank  house,  185 
Dore  furnace,  action  in,  110 

capacity  of,  192 

flue  dust,  29,  112 

metal  losses,  12 

slags,  29,  112 
Ductility,  of  refined  copper,  152 

tests  for,  152 
Dust,  (see  flue  dust) 


Efficiency,  current,  71,  79 

of  engines,  174 

of  generators,  179 

of  turbines,  175 

of  waste  heat  boilers,  172 

thermal,  138 
Electrodes,  age  of,  63 

current  leakage  between,  72 

forms  of,  46 

gassing  at,  76 

spacing  of,  51,  185 
Electrolyte,  circulation  of,  49 

composition  of,  52,  88 

conductivity  of,  51 

copj>er  in,  26,  52 

current  leakage  through,  72 

energy  required  for  heating,  57, 
180 


Filtration,  of  nickel  sludge,  99 

Flapping,  128 

Float  slime,  composition  of,  86 

in  refining  furnace,  138 
Flotation,  of  anode  slimes,  120 
Flue  dust,  dore"  furnace,  29,  112 

electrostatic  recovery  of,  12 

losses,  12 

Fuel,  (see  coal,  oil,  etc.) 
Furnace  (see  anode  furnace,  etc.) 

acid,  163 

basic,  164 

capacity,  191 

construction,  162 

fuel  consumption,  172,  191 

slag  produced  by,  11,  192 
Furnace  bottoms,  acid,  163 

anode  furnace,  35 

basic,  164 

cupola,  36 

dore",  36 

impregnation  of,  163 

refining  furnace,  35 

sampling,  161 

treatment  of,  161 


Gas  engines,  174 
Generators,  efficiency  of,  179 

electrolytic,  177 

unipolar,  179 
Gold,  effect  of,  on  conductivity,  144 

in  anodes,  85 

in  anode  furnace  slags,  137 

in  black  copper,  166 

in  bluestone,  96 


INDEX 


207 


Gold,  in  cathodes,  13,  65 
in  cathode  silver,  115 
in  cement  copper,  91,  165 
in  cupel  slag,  111 
in  Korean  coins,  166 
in  refining  furnace  slags,  137 
in  secondary  copper,  166 
in  slimes,  87,  106 
in  wirebars,  85 
refining,  115,  116,  196 
sand,  115 


H 


Heating,  electrolyte,  54,  180 

feedwater,  181 

of  contacts,  48 

of  switchboard  instruments,  188 
Hydrochloric  acid,  in  electrolyte, 
83,  88,  196 


Impurities,     concentration     of,     in 
slimes,  87 

deposition  of,  76 

distribution  of,  85 

exits  for,  84 

in  anodes,  85,  87 

in  electrolyte,  88 

in  foreign  pig  copper,  81 

in  pig  copper,  80 

in  secondary  pig  copper,  81 

in  wirebars,  85 

sources  of,  83 
Insoluble,  (see  silica) 
Insoluble  anode  tanks,  (see  liberat- 
ing tank) 

Interest,    on    metals    tied    up,    19 
Iron,  consumption  of,  in  producing 
cement,  91 

effect  of,  on  conductivity,  144 

electrolytic,  195 

in  anodes,  85,  87 

in    anode    furnace    slags,    137 

in     blister     copper,     82,      166 
'  in  bluestone,  96 

in  cement  copper,  91,  165 

in  crude  nickel  salt,  100 

in  cupel  slag,  111 


Iron,  in  dore,  112 

in  dore  furnace  slags,  1 12 

in  electrolyte,  88 

in  electrolytic  nickel,  104 

in  float  slime,  86 

in  foreign  pig  copper,  81,  166 

in  Korean  coins,  166 

in  mother  liquor  salt,  96 

in  nickel  salts,  105 

in  pig  copper,  80,  166 

in  refining  furnace  slags,  137 

in  secondary  pig  copper,  81,  166 

in  slimes,  87,  106 

in  wirebars,  85 


Junk,  sampling  of,  159 
treatment  of,  161 

K 

Korean  coins,  composition  of,   166 


Labor  cost,  at  furnaces,  141 

in  charging,  126 

relation  of,  to  current  density, 

141 
Lake    copper,    arsenical,    151,    165 

conductivity  of,  143 

premium  on,  82 

removal  of  arsenic  from,  165 

treatment  of,  124,  194, 
Leaching,  ores,  194,  195 

slimes,  110 
Lead,  effect  of,  on  conductivity,  144 

electrolytic,  195,  196,  198 

in  anodes,  85,  87 

in  blister  copper,  82,  166 

in  cupeling,  104,  111 

in  cupel  slag,  111 

in  dore  furnace  slags,  112 

in  foreign  pig  copper,  81,  166 

in  Korean  coins,  166 

in  nickel  salts,  105 

in  pig  copper,  80,  166 

in  secondary  pig  copper,  81,  166 

in  slimes,  87,  106,  112 

in  wirebars,  85 


208 


INDEX 


Lead,  losses  in  cupeling,  111 
parkes  process,  195 
pattisonizing,  195 
recovery  of,  170 

Liberating  tanks,  efficiency,  97 
proportion  of,  27,  59 
purifying  by  means  of,  97 
voltage  at,  186 

Lime,  in  anode  furnace  slags,    137 
in  cupola  slags,  11 
in  dore  furnace  slags,  112 
in  refining  furnace  slags,  137 

Location,  of  plant,  183 


M 


Magnesia,  in  boneash,  135 

in  dore  furnace  slags,  112 

Magnesite,    brick    in    furnace 

construction,  164 
Magnesium  sulphate,  in  electrolyte, 

88 
Matte,  in  dore  furnace,  112 

in  making  shot  copper,  93 
Melting,  copper,  128 
Metal  losses,  by  leakage,  15 

by  theft,  14 

by  wind,  14 

copper,  17 

from  anode  furnace  stacks,  12 

from  cupola  stack,  13 

from  refining  furnace  stacks,  12 

from  silver  refinery  stack,  12 

in  assaying,  8 

in  by-products,  13 

in  cathodes,  13 

in  cupola  slags,  12 

in  fine  gold  and  silver,  13 

in  soil,  15 

gold,  17 

silver,  17 

sources,  2 
Metal  margins,  14 
Metals  tied  up,  balancing,  31 

circulating,  23,  30 

in  anode  furnaces,  24 

in  conductors,  36 

in  furnace  bottoms,  34 

in  process,  21 


Metals,  in  refining  furnaces,  28 

in  silver  refinery,  29 

in  tank  house,  25 

relation  of,  to  current  density, 

66 

Moebius  process,  116 
Mother  liquors,  salts  produced  from, 
96 

treatment  of,  96 
Moulds,  defects  in,  136 

manufacture  of,  135 

metal  in,  37 

wash  for,  136 

N 

Nickel,  behaviour  of,  in  reverbera- 
tory,  169 

composition  of  crude  salt,  100 

composition  of  nickel  salts,  105 

electrolytic,  195 

in  anodes,  85,  87,  168 

in  anode  furnace  slags,  137 

in  blister  copper,  82,  166 

in  bluestone,  96 

in  dore,  112     „ 

in  dore  furnace  slags,  112 

in  electrolyte,  88,  89 

in  foreign  pig  copper,  81,  166 

in  mother  liquor  salt,  96 

in   secondary   pig   copper,    81, 
166 

in  refining    furnace  slags,    137 

in  slimes,  87,  106,  112 

in  wirebars,  85,  89 

market  requirements  for,  104 

offsetting  corrosion,  59. 

recovery  of,  99,  168 

uses  for,  104 
Nitric  acid,  gold  boiling,  115 

in  silver  electrolyte,  115,  196 

losses,  196 

O 

Operating  cost,  relation  to   current 
density,  68 

relation  to  size  of  plant,  184 
Oxidizing,  impurities,  137,  165 

slimes   110,  120 


INDEX 


209 


Palladium,    in    anode    slimes,     106 

recovery  of,  116 

uses  of,  101 
Parting  composition  of  gold  mud,  30 

electrolyte,  34 

electrolytic,  115,  192 
Phosphorus,    as   deoxidizing   agent, 
151 

effect  on  conductivity   of,   144 
Pig  sizes,  6 
Pitch,  defective,  150 

definition  of,  148 

determination,  148 
Platinum,  in  anode  slimes,  106 

marketed  as,  104 

recovery  of,  116 
Poling,  effect  of,  148 

overpoling,  150 

with  oil,  131 

with  wood,  131 
Power,  electrolytic,  173 

gas,  174 

house,  187 

hydroelectric,  174 

problem,  171 

relation  to  current  density,  60 

steam,  174,  175 

Purification  of  electrolyte,  by  blue- 
stone  manufacture,  93 

by  cementation,  94 

by  liberating  tanks,  97 

cylical,  98 


R 


Refining,  efficiency  of,  85 

silver,  192 

time  required  for,  19 
Refining  furnace,  charging,  126 

charging  crane,  127 

coking,  130 

continuous  melting,  141 

cycle,  136 

flapping,  128 

flue  dust,  28 

function  of,  125 

melting,  128 

14 


Refining  furnace,  poling,  131 

pouring,  133 

recharging,  127 

skimming,  130 

slag,  28,  137 

slag  formation,  136 
Roasting,  slimes,  109,  120 

S 

Sampling,  errors  in  method,  5 
furnace  charges,  128 
moisture  in,  4 

of  amalgamation  plates,  160 
of  furnace  bottoms,  161 
of  junk,  159 
of  scrap,  160 
of  slimes,  192 
precautions  in,  6 
salting,  8 
templates  in,  6 
Scrap,  (see  anode  scrap) 
sampling  of,  160 
treatment  of,  162 
Screening,  bluestone,  94 

slimes,  108,  117 
Selenium,  impurities  in,  114 
in  anodes,  85,  87 
in  blister  copper,  166 
in  dore,  112 

in  dore  furnace  flue  dust,  112 
in  dore  furnace  slags,  112 
in  Korean  coins,  166 
in  pig  copper,  166 
in  secondary  copper,  166 
in  slimes,  87,  106,  112 
in  wirebars,  85 
marketed  as,  105 
recovery  of,  113 
uses  of,  102 
Set,  defective,  150 
definition  of,  148 
determination  of,  148 
Shot  copper,  production  of,  93 
Shot  towers,  capacity  of,  193 

operation  of,  93 
Silica,  in  anodes,  87 

in  anode  furnace  slags,  137 
in  blister  copper,  166 


210  INDEX 

Silica,  in  boneash,  135  Steam,  engines,  174 

in  cement  copper,  165  exhaust  turbines,  175 

in  cupola  slags,  11  for  boiling  tanks,  181 

in  cupel  slags,  111  for  engines,  174 

in  dore  furnace  slags,  112  for  heating  electrolyte,  180 

in  Korean  coins,  166  for  heating  feed  water,  181 

in  refining  furnace  slags,  137  .         for  miscellaneous  demand,  181 

in  secondary  copper,  166  turbines,  175 

in  slimes,  106,  112  Sulphur,  effect  of,  on  conductivity, 

Silicon,  effect  of,  on  conductivity,  144 

144  in  anodes,  85,  87 

Silver,  cathode,  115  in  anode  furnace  slags,  137 

effect  of,  on  conductivity,  144              in  blister  copper,  82,  166 

impurities  in,  103  in  dore  furnace  slags,  112 

in  anodes,  85  in  electrolytic  nickel,  104 

in  anode  furnace  slags,  137  in  foreign  pig  copper,  81,  166 

in  black  copper,  166  in  Korean  coins,  166 

in  blister  copper,  166  in  pig  copper,  80,  166 

in  bluestone,  96  in  refining  furnace  slags,  137 

in  cement  copper,  91,  165  in  secondary  pig  copper,  81,  166 

in  cupel  slag,  111  in  slimes,  87,  106,  112 

in  dore,  112  in  wirebars,  85 

in  dore  furnace  flue  dust,  112  Sulphuric  acid,   in  electrolyte,    88, 

in  dore  furnace  slags,  112  196 

in  float  slime,  86  losses,  15 

in  Korean  coins,  166  parting,  114,  192 

in  mother  liquor  slat,  96  recovery  of,  99 

in  refinery  furnace  slags,  177  Switchboard,  design  of,  187 

in  secondary  copper,  166  instruments,  187 
in  slimes,  87,  106,  112 

in  wirebars,  85  T 
refinery,  192 

refining  time,  22  Tank  resistance,  analysis  of,  42 

sponge,  115  classification  of,  38 

Skimming,  130  contacts,  42 

Slags,  amount  of,  11,  192  counter  electromotive  force,  41, 

anode  furnace,  25,  137  45 

cupola,  11  electrolyte,    52 

dore"  furnace,  28,  137  transfer  resistance,  49,  55 

formation,  136  Tanks,  arrangement  of,  46 

losses,  12  asphalted,  197 

metallics  in,  137  boiling,  94,  98,  106 

refining  furnace,  28,  137  crystallizing,  94 

retreatment  of,  138,  167,  169  design  of,  185 

Slimes,  (see  anode  slimes)  Walker  system,  41 

Sodium  salts,  in  dore  furnace,  112      Tellurium,  effect  of,  on  conductivity, 

in  electrolyte,  193  144 

Starting  sheets,  age  of,  26  in  anodes,  85,  87 

proportion  required,  26  in  blister  copper,  166 


INDEX 


211 


Tellurium,  in  dore",  112 

in  dore  furnace  flue  dust,  112 

in  dore  furnace  slags,  112 

in  pig  copper,  166 

in  secondary  copper,  166 

in  slimes,  87,  106,  112 

in  wirebars,  85 

marketed  as,  106 
Temperature,  at  switchboard,  187 

in  furnaces,  163,  173 
Temperature  of  electrolyte,  effect  of, 
on  corrosion,  58 

effect  of,  on  deposit,  59 

effect  of,  on  resistance,  53 

in  practice,  60,  196 
Tensile    strength,    tests    of    copper 

wire,  153 

Thermal  efficiency,  138 
Time  required,  for  casting.  21 

for  electrolysis,  21 

for  handling,  21 

for  parting,  22 

for  refining,  19 

for  sampling,  21 

for  treating  slimes,  22 

for  weighing,  21 
Tin,  effect  of,  on  conductivity,  144 

in  Korean  coins,  166 

in  secondary  copper,  166 

recovery  of,  167,  195 
Torsion,  tests  of  copper  wire,  153 
Turbines,  exhaust,  175 

steam,  175 

U 

Unipolar  generators,  179 

V 

Vacuum,  in  condensers,  177 
Voltage,  busbar  drop,  40 
limiting,  40 


Voltage,  liberating  tank,  186 

stripping  tank,  186 

tank,  186 
Voltmeters,  precision  of,  187 

proper  location  of,  190 

W 

Walker  tank  system,  41 

Walker  wheel,  133 

Waste  heat  boilers,  size  of,  172 

steaming  rate  of,  172 

recovery  from,  138,  172 
Water,  condensing,  177 

distribution  of,  171 
Weighing,  copper,  3 

drafts  in,  3 

moisture  correction  in,  4" 

precision  of,  3 

silver  and  gold,  4 

standards  in,  3 

taring  cars  in,  4 

types  of  scales  for,  3 
Whitehead  parting  process,  116 
Wirebars,  defects  in,  149,  159 

dimensions  of,  155 

impurities  in,  85 
Wohlwill  process,  116 


Zinc,  effect  of,  on  conductivity,  144 
electrolytic,  195,  198 
in  anodes,  87 
in  blister  copper,  166 
in  dore",  112 

in  dore"  furnace  slags,  112 
in  electrolyte,  88 
in  nickel  salts,  105 
in  Korean  coins,  166 
in  secondary  copper,  166 
in  slimes,  87,  106,  112 
recovery  of,  168 


VC  33898 


338 


7W-7I 


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